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IC 


8973 



Bureau of Mines Information Circular/1984 



Mine Ground Control 

Proceedings: Bureau of Mines Technology Transfer 
Seminars, Pittsburgh, PA, December 6-7, 1983, 
and Denver, CO, December 8-9, 1983 



Compiled by Staff, Bureau of Mines 




UNITED STATES DEPARTMENT OF THE INTERIOR 






Information Circular 8973 

V 



Mine Ground Control 

Proceedings: Bureau of Mines Technology Transfer 
Seminars, Pittsburgh, PA, December 6-7, 1983, 
and Denver, CO, December 8-9, 1983 



Compiled by Staff, Bureau of Mines 




UNITED STATES DEPARTMENT OF THE INTERIOR 
William P. Clark, Secretary 

BUREAU OF MINES 
Robert C. Norton, Director 







Library of Corigress Cataloging In Publication Data: 



Bureau of Mines Technology Transfer Seminars (1983 : 
Pittsburgh, PA, and Denver, CO) 

Mine ground control. 

(Information circular / Bureau of Mines ; 8973) 

Includes bibliographies. 

Supt. of Docs, no.: I 28.27:8973. 

1. Ground control (Mining)— Congresses. I. United States. Bureau 
of Mines. II. Title. III. Series: Information circular (United States. 
Bureau of Mines) ; 8973. 

TN295.U4 [TN288] 622s [622'.28] 83-600351 



^ PREFACE 

N This Information Circular summarizes recent Bureau of Mines research 
fv^ results concerning ground control in mining environments. The papers 

represent only a sample of the Bureau's total research effort in this 

area, but they outline major portions of the program. 

:^^ The papers reproduced here were presented at Technology Transfer Sem- 

'^ inars on Mine Ground Control given in December 1983 in Pittsburgh, PA, 

xA| and Denver, CO. Those desiring more information on the Bureau's ground 

X control program or health and safety technology programs in general, or 

V information on specific situations, should contact the Bureau of Mines, 

^ Division of Health and Safety Technology, 2401 E Street, NW. , Washing- 
ton, DC 20241, or the appropriate author listed in these proceedings. 



1^ 






iii 



CONTENTS 

Page 

Preface 1 

Abstract 1 

Introduction, by Chi-shing Wang and John M. Karhnak 2 

Mine Design 

Geologic Studies in Ground Control, by Noel N. Moebs 4 

Coal and Rock Properties for Premine Planning and Mine Design, by 

Richard E. Thill 15 

Pillar Design Equations for Coal Extraction, by Clarence 0. Babcock 36 

Underhand Cut-and-Fill Stoping for Rock Burst Control, by F. Michael Jenkins 

and K. Robert Dorman 49 

Hazard Detection 

Acoustic Cross-Borehole System for Hazard Detection, by Karen S. Radcliffe and 

Richard E. Thill 64 

In-Seam Geophysical Techniques for Coal Mine Hazard Detection, by 

Richard J. Leckenby and James J. Snodgrass... 75 

Satellite Imagery as an Aid to Mine Hazard Detection, by Robert A, Speirer and 

Stanton H. Moll 89 

Microseismic Techniques Applied to Failure Warning in Mines , by 

Fred W. Leighton 96 

Mechanical and Ultrasonic Closure Rate Measurements, by Roger McVey 108 

Ground Installation Equipment 

Remote Manual Roof Bolters, by John E. Bevan 117 

Field Test of an Automated Temporary Roof Support (ATRS) Used on a Low-Coal, 
Single, Fixed-Head Roof Bolting Machine (Squirmer), by Charles T. Chislaghi 
and Thomas E. Marshall 122 

Roof Support Systems 

Development of Epoxy Grouts and Pumpable Bolts, by Robert R. Thompson 126 

Development of Lightweight Hydraulic Supports , by John P. Dunf ord 129 

Mobile Roof Support and Applications in Retreat Mining, by Robert R. Thompson.. 133 

Inorganic Grouts for Roof Bolting, by Jack E. Fraley 138 

Research, Development, and Use of Steel-Fiber-Reinforced Concrete Cribbing for 

Mine Roof Support, by Thomas W. Smelser and Dale A. Didcoct 146 





UNIT OF MEASURE ABBREVIATIONS 


USED IN 


THESE PAPERS 


A 


ampere 


lb 


pound 


cfm 


cubic foot per minute 


L/min 


liter per minute 


cm 


centimeter 


L/s 


liter per second 


deg 


degree 


m 


meter 


ft 


foot 


MHz 


megahertz 


ft/min 


foot per minute 


mi 


mile 


ft/s 


foot per second 


min 


minute 


gal 


gallon 


pin/in 


microinch per inch 


g/cm^ 


gram per cubic centimeter 


mm 


millimeter 


GPa 


gigapascal 


pm 


micrometer 


gpm 


gallon per minute 


MPa 


megapascal 


h 


hour 


ms 


millisecond 


hp 


horsepower 


US 


microsecond 


Hz 


hertz 


mV 


millivolt 


in 


inch 


ns 


nanosecond 


in/min 


inch per minute 


pet 


percent 


kHz 


kilohertz 


psi 


pound per square inch 


km 


kilometer 


s 


second 


km/s 


kilometer per second 


V 


volt 


kV 


kilovolt 


wt pet 


weight percent 


L 


liter 


yr 


year 



MINE GROUND CONTROL 

Proceedings: Bureau of Mines Technology Transfer Seminars, Pittsburgfi, PA, 
December 6-7, 1983, and Denver, CO, December 8-9, 1983 

Compiled by Staff, Bureau of Mines 



ABSTRACT 



These proceedings consist of papers presented at Bureau of Mines Tech- 
nology Transfer Seminars in December 1983 for the purpose of disseminat- 
ing recent advances in mining technology in the area of mine ground con- 
trol. The Bureau of Mines conducts several of these seminars each year 
in order to bring the latest results of Bureau research to the attention 
of the mining industry as quickly as possible. 



INTRODUCTION 
By Chi-shing Wang "I and John M. Karhnak'' 



The basic goal of the Bureau of Mines 
Ground Control research program is to 
provide the mining industry with technol- 
ogy that will lead to the reduction of 
accidents due to falls of ground. The 
problems of ground control are the ina- 
bility to "see" geologic anomalies ahead 
of the mine workings, the difficulty 
in predicting ground movements induced 
by excavation, and the need to provide 
efficient ground control over the wide- 
ly varying, and frequently unexpected, 
conditions encountered from one place 
to another. Current program objectives 
include — 

Exploring new methodologies and vali- 
dating available tools for predicting 
hazardous geologic features in ad- 
vance of mining and for selecting and 
designing safe and efficient mining 
systems and layout plans. 

Developing Instrumentation methods 
and equipment able to detect and warn 
of imminent failure of ground during 
mining. 

Ascertaining the most effective per- 
manent support methods for specific 
ground conditions. 

Investigating roof support installa- 
tion methods that improve working 
conditions at the face and remove ma- 
chine operators from dangerous unsup- 
ported areas. 

Advance knowledge of potentially haz- 
ardous geologic features is a prerequi- 
site to success in designing out the as- 
sociated safety risks from the mining 
plan during the early stage of mine de- 
sign and development. To enable mine op- 
erators to satisfy this requirement, the 
Bureau has continued its efforts in 

^ Staff engineer, Division of Health and 
Safety Technology, Bureau of Mines, Wash- 
ington, DC. 



assessing the applicability of various 
geologic mapping methods and geophysical 
techniques to detection and delineation 
of geologic anomalies and mine voids from 
the surface, boreholes, or mine workings. 

The ability of miners to predict poten- 
tial hazards ahead of working faces and 
to detect impending ground failures at 
the working place during mining is cru- 
cial to timely evacuation and safety pre- 
cautions to avoid injuries. Therefore, 
Bureau research continues to assess in- 
seam geophysical techniques for detecting 
geologic anomalies and mine voids ahead 
of coal faces in underground coal mines , 
automatic microseismic roof fall warning 
systems, and mechanical and ultrasonic 
devices for continuous measurement of 
roof-to-floor closure rates. 

Better understanding and prediction 
capability of mine geology and ground 
movements serve as preventive measures 
permitting mine operators to design out 
potential ground hazards, and enabling 
miners to avoid or escape from dangerous 
areas in time. However, protective mea- 
sures for assurance of safe mine workings 
with an adequate ground support are es- 
sential to maintaining active mining op- 
erations. Hence, the Bureau has invested 
considerably in research of more cost- 
effective and more efficient temporary 
and permanent roof support systems, al- 
ternative ground support materials, and 
remotely operated roof bolt installation 
equipment and procedures. 

This technology transfer seminar intro- 
duces to the mining industry the signifi- 
cant technological advances made by Bu- 
reau research in recent years in various 
aspects of ground control in mines. Em- 
phasis is placed on those innovations 
that have been validated by industrial 
applications or by extensive field demon- 
strations. Although many of those stud- 
ies and innovations were conducted in or 
designed for use in coal mine ground 



control, most of the new technologies 
can be adapted for use in various mining 
environments including metal and nonmetal 
mines. As the current research policy 
of the U.S. Government is to shift empha- 
sis toward long-term, high-risk, and 
high-payoff basic research, the mining 



industry and mining equipment manufac- 
turers are encouraged to assess the po- 
tential of immediate uses or private 
investments in commercial development of 
Bureau research products that are offered 
at this seminar. 



MINE DESIGN 

GEOLOGIC STUDIES IN GROUND CONTROL 
By Noel N. Moebs' 



ABSTRACT 



Several geologic techniques are re- 
viewed that are useful in planning mines 
and in solving ground control problems. 
A summary is presented of a study of coal 



mine roof structures and the use of 
this information in improving supple- 
mentary support and anticipating support 
problems. 



INTRODUCTION 



The Bureau of Mines has been conducting 
studies directed toward increasing the 
utilization of geologic methods in coal 
mine ground control, and more recently in 
quarry operations, for the purpose of re- 
ducing hazards from roof and rock falls. 
It is commonly recognized that geology is 
the key to effective ground control; that 
is , a sound knowledge of the character 
and structure of rock provides a sound 
basis for mine planning and selection of 
appropriate roof support methods. 

For example, studies by the Bureau 
of Mines have identified geologic struc- 
tures in mine roof rock that contrib- 
ute to many roof falls in Appalachian 
coal mines. These structures, includ- 
ing paleochannels , kettlebottoms , scours, 
pinchouts, slickensides , clay veins, cre- 
vasse splays, and joints, can often be 
identified during, and sometimes before, 
mine development. Mine projections can 
be revised to reduce the adverse effects 
of discontinuities in roof structure, 
large roof areas of laminated sandstone 
or incompetent strata generally can be 
delineated or inferred from exploratory 
drill-hole data, and the need for sup- 
plementary support can be anticipated. 
Accurate descriptions of roof geology 
also provide some indication of optimum 
length and type of roof bolts that should 
be installed. 

^Geologist, Pittsburgh Research Center, 
Bureau of Mines, Pittsburgh, PA. 



In addition to studying geologic struc- 
tures in coal mine roof , which are de- 
scribed in another section, the Bureau 
investigated the concept of a hazard map 
that would delineate potentially hazard- 
ous zones prior to mining. A method of 
preparing a hazard map through the compu- 
ter processing of drill-log data was de- 
veloped, and the method was demonstrated 
by preparing a hazard map of a small por- 
tion of the Pittsburgh Coal Basin. This 
methodology also should be applicable to 
individual mines where an adequate den- 
sity of exploratory drill holes is avail- 
able. A computer program is available 
for drill-hole data processing, which re- 
sults in a printout map of anticipated 
mining conditions. The identification 
and weighting of hazardous geologic vari- 
ables are provided for by modification of 
the program. 

Another geologic tool applied to ground 
control studies by the Bureau is linear 
analysis, 2 whereby an attempt is made to 
correlate linear features from aerial 
photographs or satellite images with sub- 
surface rock stability conditions. The 
results of this research in the Appa- 
lachian coal region have, to date, been 
inconclusive in that at some locations a 

^Linear, or photolineament, is a line 
on an aerial photograph (or transposed 
from satellite imagery) that is con- 
trolled by an alignment of surface or 
near-surface structures or conditions. 



correlation was discovered but at many it 
was not. The lack of precise data on the 
location of roof failure, the failure to 
identify the "ground truth" or true geo- 
logic character of a linear, and the 
inability to selectively choose valid 
lineaments all have severely limited the 
practical application of linear analysis 
to ground control studies. Continued re- 
search is urgently needed to establish 
the limitations and the potential of 
linear analysis as a valid geologic meth- 
od in ground control studies. 

Virtually the only direct method of ob- 
taining data on subsurface strata prior 
to mining is through core drilling. Un- 
fortunately, most drill core in the coal 
regions has been logged by the driller 
himself , and while this might have been 
adequate to determine coal thickness and 
general rock units, it is entirely inade- 
quate for use in the assessment of an- 
ticipated roof conditions, cavability, or 
paleoenvironmental reconstructions to 
determine coal continuity and quality. 
Even with the increasing use of trained 
professional geologists to log drill 
core, there is still much to be desired 
in the accurate and uniform identifica- 
tion of rock types. Recognizing that 
ground control investigations prior to 
mining require the best possible drill- 
core records , especially where no provi- 
sion is made to preserve the core, the 
Bureau has developed field guides for the 
identification of cored rock in the 
Pittsburgh Basin of Pennsylvania, Ohio, 
and West Virginia, and in the Pocahontas 
Basin of southeastern West Virginia. 
These guides are pocket size, on weather- 
proof paper, with a printed key to rock 
types and color photographs of the actual 
core. The Bureau produced these guide- 
books on an experimental basis. They 
have proven to be immensely popular and 
useful, and as they become commercially 
available a great deal more information 
should be extracted from drill core than 
before. 



While most geologic studies have been 
directed toward improving coal mine 
ground control, a project recently initi- 
ated in slate quarry operations utilizes 
some of the same research methodologies 
and thus is included in this overview. 
The ground control problems of slate 
quarry operations result partly from 
geologic discontinuities in the vertical 
quarry walls , which range from 80 to 560 
ft in height (fig. 1). These geologic 
discontinuities account for an occasional 
rockfall or sudden collapse of an entire 
highwall. These failures occur despite 
scaling and pinning. Bolts are not em- 
ployed for highwall support. Some rock 
pressure problems are evident on the 
quarry floors; the integrity of rock 
barriers separating quarries sometimes is 
in doubt ; and during the winter months 
ice accumulations at the quarry brinks 
are severe, and large masses or ice 
stalactites that become dislodged consti- 
tute a real hazard. Research to abate 
these ground control problems includes 
the use of strain gauges to detect immi- 
nent falls of rock, the identification 
and mapping of geologic features that 
have contributed to highwall collapse, 
the measurement of developed rock stress- 
es, and the testing of both mechanical 
and resin bolts to secure wall rock. 

Following this brief review of some 
geologic studies for ground control, the 
remainder of this paper will summarize 
the results of research on hazardous geo- 
logic structures in coal mine roof. This 
research provides an example of how the 
recognition of coal mine roof structures 
can lead to improved mine planning and 
supplementary support. A bibliography of 
selected references on this topic is pro- 
vided at the end of this paper. Because 
of the wide variance and diversity of 
geologic structures in mine overburden, 
only the most significant ones are dis- 
cussed in the following sections. 




FIGURE 1, - Highwall of slate quarry operation. 



COAL MINE ROOF STRUCTURES 



Among the causes of coal mine roof 
failure are the hazardous roof conditions 
produced by certain geologic structures. 
The role of these structures in mine roof 
falls is not always apparent because the 
formations involved are sometimes poorly 
exposed or difficult to recognize without 
special training and underground experi- 
ence. Most of the structures are small 
scale, less than a few tens of feet in 
width. In the underground coal produc- 
tion environment, these structures com- 
monly are unrecognized or are ignored by 
all except those responsible for the 
proper roof support installation. 

PALEOCHANNELS 

A paleochannel, or miners' "roll," is 
the trough-shaped remnant of an ancient 
stream channel that has been cut into 
older rock, such as the roof shale and 
coalbed. The channel subsequently has 
been filled in with younger sediments, 
usually sandstone. The paleochannels 
most troublesome in mine roof support 
seldom exceed about 30 ft (9.3 m) in 
width and range from a few tens to a few 
hundreds of feet in length. 

Channels that are more than 30 to 50 ft 
(9.3 to 15.5 m) wide and have cut down 
through most of the coalbed are termed 
"washouts." Although washouts can be a 
severe problem in mine development or 
longwall operations , they generally do 
not cause serious roof fall problems for 
more than about 50 ft (15.5 m) outside 
the margins of the channel. 

The most typical paleochannel is a 
linear feature, often extending across 
several adjacent entries. The slicken- 
sides on the undersurf aces of the U- 
shaped channel constitute planes of weak- 
ness and a lack of bonding. Thus the 
weak shales and coal adjacent to and 
directly beneath the sandstone-filled 
channel separate readily from it as soon 
as the underlying coalbed is mined. The 
most adverse roof condition occurs when 
the channel trend is parallel to an en- 
try, resulting in virtually continuous 



roof support problems. Some shale-filled 
channels also exist and constitute a sim- 
ilar problem. 

Paleochannels are best recognized by 
their troughlike shape, slickensided mar- 
gins, "horse-belly" appearance, and gen- 
erally hard, crossbedded, poorly sorted 
sandstone filling. Methods that have 
been most often successful in supporting 
shale channel margins are strapping and 
the use of wood headers, and, for severe 
cases, posts and crossbars. The use of 
angle bolts that anchor into the hard 
channel filling (fig. 2) is suggested as 
another possible means of support. Un- 
derground mapping is the best method of 
determining the trends of channels. 
Sometimes mine intersections can be re- 
located to avoid an individual channel, 
but, more often, several parallel chan- 
nels occur, and the only option is to re- 
vise the entry projections to intersect 
channels at an obtuse angle. 

Paleochannels in roof strata can rarely 
be detected in advance of mining by nor- 
mally spaced exploratory core drilling 
(approximately 2,500-ft centers) because 
of their small dimensions. However, 
their presence can be inferred in areas 
where drill core data indicate that 




- UnderclQy 



FIGURE 2. - Useof angleboltstosupportshale 
strata at channel margins. Dashed line outlines 
rock not fully supported by vertical bolting. 



thick, lenticular, crossbedded sandstone 
occurs close to the top of the coalbed. 

KETTLEBOTTOMS 

A kettlebottom is a columnar mass of 
rock embedded in and comprising a part of 
the coal mine roof strata. Kettlebottoms 
are the preserved casts of ancient tree 
stumps that grew in the peat swamp, now 
the coalbed. Once undermined, unsupport- 
ed kettlebottoms can detach from the roof 
at any time without warning, presenting a 
hazard to miners. The size and frequency 
of kettlebottoms in mine roof is depen- 
dent upon geologic events and biological 
processes active during deposition of 
roof sediments. The preserved casts tend 
to be small local features of erratic oc- 
currence, which cannot be detected by 
core drilling. They occur in the coal 
measures throughout the Appalachians , but 
are most abundant in the Pottsville age 
deposits of southern West Virginia and 
eastern Kentucky. All kettlebottoms not 
dislodged after initial mining should be 
properly supported. 

SCOURS 

The erosional action of an ancient 
stream produces an almost endless variety 
of cut-and-fill structures in rock. 
Scours, high curved, oblong, or saucer- 
shaped channel structures, must be ap- 
proached with great caution, and the same 
roof support methods should be utilized 
as with the linear paleochannels. 

Scours cannot be projected from entry 
to entry as can the persistent and linear 
channels, although, as with channels, 
their presence in mine roof is more prob- 
able where drilling indicates thick, len- 
ticular, crossbedded sandstone close to 
the top of the coalbed. 

PINCHOUTS 

The term "pinchout" is used here to 
designate the abrupt termination of 
a roof stratum (fig. 3). Some pinch- 
outs have been observed near the flanks 
of channels and washouts, and presuma- 
bly were formed by the same cutting 



action. Others appear to be due to very 
rapid thinning during normal sediment 
deposition. 

If a pinchout occurs in a stratum that 
is fairly thick (more than 1.5 ft or 0.5 
m) , competent, and located in the immedi- 
ate roof, the beam strength of the roof 
will be weakened seriously by the discon- 
tinuity. Pinchouts are not easily de- 
tected until exposed by a roof fall. 
They can sometimes be discovered or in- 
ferred during drilling of roof bolt holes 
when the roof bolter finds a pronounced 
contrast in penetration rates between op- 
posite sides of the entry. A pinchout 
requires effective roof bolting to re- 
inforce the weakened beam structure in 
the roof through the normal roof support 
plan or supplementary longer bolts that 
anchor into competent overlying strata. 

The pinchout most troublesome in mine 
roof is a relatively small local feature 
and cannot be detected by normal explora- 
tory core drilling. However, its pres- 
ence can be inferred in areas where the 
same individual sandstone strata close to 
the coalbed do not occur in adjacent 
holes. 

SLICKENSIDES 

Probably the most common hazardous roof 
structure occurring throughout the Appa- 
lachian coal region is the slickenside, 
or "slip." It is best developed and most 
common in highly argillaceous rocks such 
as shales, claystones, or miners' "clod." 
A slickenside is a smooth, polished, and 
sometimes striated or grooved surface re- 
sulting from movement of rock on either 
side of the surface. A slickenside usu- 
ally is curved and in coal mine roof rock 
generally is convex toward the coalbed. 

The slickenside constitutes a disconti- 
nuity in the beam roof structure and 
therefore weakens the roof, creating a 
potential hazard unless detected and 
properly supported. It is difficult to 
estimate the extent of the slickenside 
simply by observing the trace in the mine 
roof, but it can probably be assumed that 
the longer the trace, the greater the 
extent. 




Coalbed 



Underclay- 



_4Ft 



Scale 



FIGURE 3. - Abrupt termination of sandstone roof stratum. Dashed lines show roof structure when 
first exposed; original roof was at top of the coalbed. 



Small-sized slickensldes (less than 3 
ft or 0.9 m) , nearly always found in mine 
roof , are adequately supported by the 
normal roof support plan at each mine. 
However, large slickensldes should be re- 
garded with extreme caution, and some 
form of supplementary support used. In 
practice, this commonly consists of ex- 
tending the support area of the roof 
bolts by using straps or wood headers , 
installing additional bolts, or angle 
bolting to more likely penetrate the 
slickenside at a right angle and avoid 
cantilevered segments of roof rock (fig. 
4). In situations where heavy concentra- 
tions of slickensldes are encountered, it 
is advisable to consider using posts and 
crossbars, full column resin bolts, or, 
in severe cases, roof trusses. 



There is currently no practical method 
of detecting slickensldes in advance of 
mining, although, if core drilling in- 
dicates that the immediate roof con- 
sists of a nonlaminated claystone, then 
slickensldes almost certainly will be 
present. 

CLAY VEINS 

Clay veins, or more properly, claystone 
dikes, are wedge-shaped masses of slick- 
ensided claystone or mudstone filling a 
crevice in a coalbed (fig. 5). These 
generally range up to 6 ft (2 m) in width 
and persist for sometimes hundreds of 
feet in length. 



10 





:.Underclay 



FIGURE 4. - Angle bolting of slickensides. 



Clay veins are particularly prevalent 
and troublesome in the Pittsburgh Coalbed 
near Wheeling, WV, but occur sporadically 
throughout the Appalachian region and in 
many other coalbeds. Sometimes the term 
"spar" is applied to a very narrow clay 
vein or one that extends downward only a 
few inches into the coalbed. Other types 
of clay veins, such as steeply dipping, 
narrow, claystone-f illed joints, occur in 
the Appalachian region but are rare and 
seldom cause roof problems. 

The weakening effect of clay veins on 
mine roof extends from 3 to 12 ft (0.9 to 
3.7 m) above the top of the coalbed. In 
addition to forming a discontinuity in 
the coalbed and immediate roof , the clay 
vein and surrounding rock are heavily 
slickensided and therefore likely to 




FIGURE 5. - Clay vein in entry rib and roof. 



11 



break away from the roof in thick blocks 
as the supporting coalbed is mined. The 
slickensides along the margins and in- 
terior of a clay vein tend to be aligned 
parallel to the trend of the vein and to 
dip toward the center of the vein forming 
a troughlike effect. While prompt bolt- 
ing, blocking, and strapping prevent im- 
mediate falls of roof, the clay-rich rock 
is susceptible to slaking and gradually 
disintegrates, falling little by little. 

Clay veins have rarely, if ever, been 
detected by exploratory core drilling be- 
cause of their narrow width. They are 
particularly abundant and should be ex- 
pected where the immediate roof consists 
of a thick, clay-rich rock, although they 
also occur immediately beneath limestone, 
sandstone, or shale strata where perhaps 
the clay-rich layer has been removed by 
erosion. Clay veins can be predicted in 



advance of mining only by underground 
mapping and judicial projection along 
their trends. 

CREVASSE SPLAYS 

The term "crevasse splay," used here in 
a nongenetic, descriptive sense, desig- 
nates a lithologic unit consisting of 
sandstone thinly interbedded with shale, 
or thin-bedded, laminated, micaceous 
sandstone (miners' "stackrock") . The 
unit ranges from 6 to 30 ft (2 to 9.3 m) 
in thickness and persists laterally as a 
sheetlike or lenticular body (fig. 6). 

The splay type deposit that is sig- 
nificant in mine roof stability is the 
so-called low splay, where a predominant- 
ly flat-bedded, laminated sandstone unit 
lies within 6 to 10 ft (2 to 3 m) of the 
top of the coalbed and, therefore, 




FIGURE 6. ■ Example of splay type deposit in mine roof. 



12 



constitutes part of the inunediate roof. 
Splays that occur much higher than 10 ft 
(3m) above the coalbed do not directly 
affect roof conditions. 

Roof falls in splay sequences generally 
occur when a separation along a bedding 
plane occurs at or above the horizon of 
the roof bolt anchors, and the bolts are 
unable to maintain the integrity of the 
rock spanning the immediate roof. Some- 
times failure can be prevented by stag- 
gering the lengths of roof bolts so 
that tensile stresses are not concen- 
trated along one horizon. Full-column 
resin bolts resist slippage along bedding 
planes and thus reduce roof sag, which 
might reduce the possibility of failure. 

A low splay deposit can be inferred 
from exploratory drill core information 
that indicates a laterally widespread, 
thin-bedded, laminated sandstone unit 
lying within 6 to 10 ft (2 to 3 m) of 
the coalbed, generally with some shale 
above and below and situated adjacent to 
channel-fill formations. 

JOINTS 

The term "joint" is used here to desig- 
nate a nearly vertical, planar fracture 
in rock, as distinguished from a slicken- 
side, which is a curved, grooved surface 
along which some movement has occurred. 
In the central Appalachian coal region, 
where coal measure strata are nearly 
horizontal, joints in mine roof rock 
usually are intraf ormational and of minor 



significance to roof stability. Roof 
falls may be terminated by a joint plane, 
but there is little evidence to indicate 
that the presence of a joint directly 
contributes to a fall. 

The situation is different along the 
eastern margin of the Appalachian coal 
region where the strata have been de- 
formed by folding and faulting. Here, 
jointing is more highly developed and 
significantly contributes to roof prob- 
lems, especially where the joints occur 
as closely spaced, parallel sets or 
groups. Highly jointed roof sometimes 
can be supported satisfactorily with 
bolted channel, strap, or mat, but severe 
conditions may require trusses or posts 
and crossbars. 

Highly jointed roof is very difficult 
to predict with certainty, but some de- 
gree of success has been claimed by peo- 
ple in the field, using interpretation 
of high-altitude aerial photography and 
satellite imagery, with the assumption 
that at least some photolineaments repre- 
sent rock joint sets. It is essential to 
conduct a critical field check to deter- 
mine the so-called ground truth of a lin- 
ear lest cultural features be mistaken 
for geologic structures. In some areas, 
linear images are quite abundant , and the 
problem is to distinguish between those 
that are significantly related to roof 
conditions or joints and those that are 
not — a task for the most skilled and con- 
scientious technologist. 



DISCUSSION 



There are a number of reasons for mine 
roof failure. The cause of some roof 
falls is obscure and attributed to poorly 
understood stress concentrations sur- 
rounding the mine openings; these falls 
occur most commonly when mine entries are 
located beneath stream valleys, and topo- 
graphic relief is 200 ft (62 m) or more. 
Some falls result from poor roof-bolting 
techniques or improper installation of 
bolts. However, most roof falls associ- 
ated with the geologic hazards referred 
to in this paper occur because unusual 



geologic structures are not recognized or 
anticipated and adequate support is not 
provided. 

The identification of the structures 
that have been described here will re- 
quire either the services of a geologist 
with underground experience or the train- 
ing of miners (especially mining machine 
and roof -bolting machine operators) and 
their supervisors. Through regular ex- 
amination of mine roof structures and 
conditions and the recording of this 



13 



information systematically on the operat- 
ing maps , and through some trial and er- 
ror, it should become possible to deter- 
mine the trends of hazardous structures 
so that mine entry projections can be re- 
vised or potentially troublesome zones 
anticipated, roof support practices up- 
graded, and roof falls reduced. 

The task will not be easy because of 
the complex character of many geologic 
features, but, in consideration of both 



the high priority of accident prevention 
and the immense financial investment at 
stake in developing a mine, the attempt 
to identify and properly support roof 
structures should become an integral part 
of every underground operation. 

The application of geologic methods to 
other mines and quarry operations should 
provide similar opportunities to improve 
ground control practices and thus reduce 
the inherent safety hazards. 



BIBLIOGRAPHY 



Alison, D. R. , E. T. Ohlsson, and K. V. 
Whitney. Geologic and Engineering Data 
Acquisition for Underground Coal Mine 
Ground Control (contract J0395010, Arthur 
D. Little, Inc.). BuMines OFR 89-80, 
1980, 98 pp.; NTIS PB 80-219272. 

Chase, F. E., and G. P. Sames. Kettle- 
bottoms: Their Relation to Mine Roof and 
Support. BuMines RI 8785, 1983, 12 pp. 

Cummings , R. A., M. M. Singh, and N. N. 
Moebs. Effect of Atmospheric Moisture on 
the Deterioration of Coal Mine Roof 
Shales. Min. Eng. (N.Y.), v. 35, No. 3, 
Mar. 1983, pp. 243-245. 

Cummings, R. A., M. M. Singh, S. E. 
Sharp, and A. W. Laurito. Control of 
Shale Roof Deterioration With Air Temper- 
ing (contract J0188028, Engineers Inter- 
national, Inc.). Volume 1 — Field and 
Laboratory Investigations. BuMines OFR 
41(l)-82, 1981, 162 pp. 

Ellenberger, J. L. Slickenside Occur- 
rence in Coal Mine Roof of the Valley 
Camp No. 3 Mine Near Wheeling, W. Va. 
BuMines RI 8365, 1979, 17 pp. 

Ferm, J. C. , R. A. Melton, G. D. Cum- 
mins, D. Mathew, L. L. McKenna, C. Muir, 
and G. E. Norris. A Study of Roof Falls 
in Underground Mines on the Pocahontas 
No. 3 Seam, Southern West Virginia and 
Southwestern Virginia (contract H0230028, 
Univ. SC). BuMines OFR 36-80, 1978, 83 
pp.; NTIS PB 80-158983. 



Hylbert, D. K. Delineation of Geologic 
Roof Hazards in Selected Coal Beds in 
Eastern Kentucky, With Landsat Imagery 
Studies in Eastern Kentucky and the 
Dunkard Basin (contract J0188002, More- 
head State Univ.). BuMines OFR 166-81, 
1980, 97 pp.; NTIS PB 82-140336. 

. Developing Geological Struc- 
tural Criteria for Predicting Unstable 
Mine Roof Rocks (contract H0133018, More- 
head State Univ.). BuMines OFR 9-78, 
1977, 246 pp.; NTIS PB 276 735. 

Jansky, J. H. , and R. F. Valane. Cor- 
relation of LANDSAT and Air Photo Linears 
With Roof Control Problems and Geologic 
Features. BuMines RI 8777, 1983, 22 pp. 

McCulloch, C. M. , P. W. Jeran, and 
C. D. Sullivan. Geologic Investigations 
of Underground Coal Mining Problems. Bu- 
Mines RI 8022, 1975, 30 pp. 

Moebs, N. N. The Geological Character 
of Some Coal Wants at the Westland Mine 
in Southwestern Pennsylvania. BuMines 
RI 8555, 1980, 25 pp. 

. Geologic Guidelines in Coal 

Mine Design. Paper in Ground-Control As- 
pects of Coal Mine Design. Proceedings: 
Bureau of Mines Technology Transfer Sem- 
inar, Lexington, KY, March 6, 1973. Bu- 
Mines IC 8630, 1974, pp. 63-69. 



14 



Moebs, N. N. Roof Rock Structures and 
Related Roof Support Problems in the 
Pittsburgh Coalbed of Southwestern Penn- 
sylvania. BuMines RI 8230, 1977, 30 pp. 

. Subsidence Over Four Room-and- 

Pillar Sections in Southwestern Pennsyl- 
vania. BuMines RI 8645, 1982, 23 pp. 

Moebs, N. N. , and E. A. Curth. Geolog- 
ic and Ground Control Aspects of an Ex- 
perimental Shortwall Operation in the Up- 
per Ohio Valley. BuMines RI 8112, 1976, 
30 pp. 

Moebs, N. N. , and J. L. Ellenberger. 
Geologic Structures in Coal Mine Roof. 
BuMines RI 8620, 1982, 16 pp. 

. Hazardous Roof Structures in 

Appalachian Coal Mines. Ch. in Ground 



Control in Room-and-Pillar Mining. AIME, 
New York, 1982, pp. 9-16. 

Moebs, N. N. , and J. C. Perm. The Re- 
lation of Geology to Mine Roof Conditions 
in the Pocahontas No. 3 Coalbed. BuMines 
IC 8864, 1982, 8 pp. 

Overbey, W. K. , Jr., C. A. Komar, and 
J. Pasini III. Predicting Probable Roof 
Fall Areas in Advance of Mining by Geo- 
logical Analysis. BuMines TPR 70, 1973, 
17 pp. 

Stingelin, R. W. , J. R. Kern, and 
S. L. Morgan. Pre-Mining Identification 
of Hazards Associated With Coal Mine Roof 
Measures (contract J0177038, HRB-Singer, 
Inc.). BuMines OFR 167-81, 1979, 206 
pp.; NTIS PB 82-140344. 



15 



COAL AND ROCK PROPERTIES FOR PREMINE PLANNING AND MINE DESIGN 
By Richard E. Thill'' 



ABSTRACT 



Nearly all phases of the mining opera- 
tion require input on the engineering 
properties of rock. Engineering proper- 
ties are particularly useful in the pre- 
mine investigation and mine design phases 
of the mining operation. 

Because of the general scarcity of en- 
gineering property data for coal measure 
strata in major U.S. coal districts and 
the lack of organized data bases to house 
and disseminate such information, the 



Bureau of Mines undertook a wide-ranging 
testing program to determine engineering 
properties of coal measure rocks in sev- 
eral U.S. coal basins. The Bureau also 
initiated the development of a data base 
for the housing, sorting, manipulation, 
and retrieval of property data by com- 
puter. Typical results are given for 
both field and laboratory determinations 
of geotechnical, mechanical, and geophys- 
ical properties. 



INTRODUCTION 



Engineering properties are important in 
nearly every phase of mining, but partic- 
ularly in mine design, layout, and ground 
control. Reviews of current literature 
revealed that (1) not many data exist for 
the engineering properties of strata as- 
sociated with U.S. coal deposits, (2) 
data available are scattered in the lit- 
erature and difficult to retrieve, and 
(3) very few data exist in data sets that 
permit correlations to be made between 
properties for predictive purposes. 

Because of the general lack of data 
in major U.S. coal-producing districts, 
the Bureau of Mines undertook a testing 
program to determine a variety of engi- 
neering properties, including sets of 



geological, geophysical, physical, and 
mechanical properties, in several major 
U.S. coalfields. In addition, data from 
earlier Bureau studies on all types of 
crystalline and sedimentary rock were 
summarized in data tables and incorpo- 
rated into a rock property data base. 
The testing program and typical examples 
of data and analyses for the coal measure 
rocks, and a brief description of the 
mechanical properties data base and data 
tables are presented in this paper. The 
property test results are being published 
in Bureau and outside publications to 
assist mine designers and planners in the 
design of openings and support struc- 
tures, and in the selection of appropri- 
ate excavation and support equipment. 



COAL MEASURE ROCK TEST PROGRAM 



The testing program consisted of both 
laboratory and field investigations on 
rock core taken at several field sites 

'Supervisory geophysicist, Twin Cities 
Research Center, Bureau of Mines, Minne- 
apolis, MN. 



(fig. 1). In the field, lithologic and 
geotechnical descriptions of core were 
made and downhole geophysical logging 
was performed. Core then was boxed 
and shipped to the Bureau for property 
testing. Portions of the tests in Illi- 
nois rock were conducted under contract. 



16 



PROGRAM ORGANIZATION 



In situ investigations 



Core 
extraction 



Lithologic 
logging 



Geotechnical 
logging 



Geophysical 
logging 



Specimen selection 
and preparation 



Laboratory investigations 



Geological 
properties 



Acoustic 
properties 



Electrical 
properties 



Mectionical 
properties 



Physical and index 
properties 



Data reduction 
and analyses 



FIGURE 1. - Cool measure rock testing program. 



In situ properties determined included 
compressional (P) and shear (S) wave 
velocities, formation density, derived 
elastic moduli, and other geotechnical 
properties. Laboratory determinations 
included mechanical properties such as 
uniaxial, triaxial, and tensile strength, 
and static elastic properties; acous- 
tic properties such as P- and S-wave 
velocities, dynamic elastic moduli, and 
acoustic impedance and attenuation; 
petrographic properties; electrical prop- 
erties such as dielectric constant, dis- 
sipation factor, conductivity, and skin 
depth; and physical or index properties 



such as bulk density, porosity, permea- 
bility, shore hardness, slake durability, 
and point-load strength index. 

MINESITES 

The several minesites investigated in- 
cluded the Gateway Mine in Pennsylvania, 
the York Canyon Mine in New Mexico, and 
three sites in the Illinois Basin. Ta- 
ble 1 indicates the types of studies un- 
dertaken at each site. The Gateway and 
York Canyon studies were more extensive 
and comprehensive than those in the 
Illinois Basin. 



17 



TABLE 1. - Properties determined at different sites 



Property test or 


Gateway Mine, 

southwestern 

Pennsylvania 


York Canyon 

Mine, northern 

New Mexico 


Illinois Basin 


characteristic 


East-central 
Illinois 


Southern Illinois 
and Indiana 


In situ: 

Lithologic 

Geotechnical 

Geophysical 

Laboratory: 

Petrographic 

Mechanical. ........ 


X 
X 
X 

X 
X 
X 
X 
X 

X 

X 


X 
X 
X 

X 
X 
X 

X 

X 

X 


X 
X 
X 

X 
X 
X 

X 

X 

X 


X 
X 


Acoustical 

Electrical 


X 


Physical and index. 

Environmental 
effects: 
Stress ............. 


X 
X 


Moisture 


X 



The Gateway Mine is located in the 
northern part of the Eastern Coal Prov- 
ince (fig. 2). Core was taken with a 



conventional drill rig equipped with a 
20-ft-long, wire line retrievable, inner 
core barrel (fig. 3). Roughly 500 ft of 




100 



Scale, mi 



200 



wv 



LEGEND Scale, mi 

Underlain by Pittsburgh Coalbed 



10 



h! '. ,' H Gateway mines 
<-\-^ Anticline 
-H^ Syncline 

FIGURE 2, - Location of Gateway minesite. 



18 




FIGURE 3. - Drill rig at the Gateway site. 



19 




FIGURE 4. - Core extraction at the Gateway site. 



core was taken at the site (fig. 4) 
and logged for lithology, percent re- 
covery and rock quality (RQD). Three- 
dimensional acoustic and gamma-gamma den- 
sity surveys also were conducted in the 
borehole. 

The York Canyon site is in the north- 
eastern portion of the Southwest Coal 
Province (fig. 5). Figure 6 shows the 
drill rig used at the site. As core was 
taken from the hole, natural moisture 
content was preserved by wrapping the 
core in tinfoil and sealing it with wax 
(fig. 7). 

The localities of the Illinois Coal 
Basin sites are shown in figure 8. 
The Danville, IL, site is located in 
the east-central portion and the other 
two sites are located in the south- 
east portion of the Interior Coal 
Province. 




FIGURE 5. - Location of York Canyon Mine. 



20 





FIGURE 7. = Sealing moisture content in the 
York Canyon core. 



INDIANA 



FIGURE 6, - Drill rig at the York Canyon site. 

LABORATORY TESTING 

Test specimens were prepared from the 
core arriving from the field. Laboratory 
property determinations included petro- 
graphic, acoustic, electrical, mechani- 
cal, physical, and index properties. 
Specimen preparation and testing was con- 
ducted in accordance with standardized 
procedures described in the Bureau of 
Mines Test Procedures for Rocks {l)^ and/ 
or as recommended by the International 

^Underlined numbers in parentheses re- " 
fer to items in the list of references at 
the end of this paper. 




FIGURE 8. - Location of Illinois sites. 



21 




FIGURE 9. - Mechanical property testing apparatus. 



Society of Rock Mechanics (2^) . The me- 
chanical property testing facilities in- 
clude two hydraulic, servo-controlled 
testing machines, one of 200,000-lb and 
the other of 500,000-lb load capacity 



(fig. 9). For triaxial confinement test- 
ing, separate servo controls are used 
for applying the uniaxial and confining 
pressure. 



RESULTS 



FIELD GEOTECHNICAL AND 
GEOPHYSICAL PROPERTIES 

As core was taken in the field, litho- 
logic descriptions were made to construct 
stratigraphic columns for the vicinity of 
the working coal seam and the overburden. 
As indicators of rock quality, percent 
recovery, RQD, and fracture frequency 
were measured. Figure 10 represents re- 
sults obtained at the Gateway minesite. 
Poor quality rock is designated by zones 
having low RQD, low core recovery, or 
high fracture frequency. 



Borehole geophysical logs were run at 
the Gateway, York Canyon, and Danville 
sites. The logs obtained at the Gateway 
site (fig. 11), for instance, permit 
comparison between elastic wave velocity 
and density response. Coal seams and 
fracture zones are readily detected by 
anomalously low-velocity, low-density re- 
sponse. From the full waveform velocity 
log, both P- and S-wave velocities can be 
determined, and when combined with den- 
sity, these permit calculation of the dy- 
namic elastic moduli, indicating stress- 
strain response and, by correlation, the 



22 



LITHOLOGIC 
LOG 
EL -7 



DRILLER'S 
LOG 
EL-7 



Or 



SO- 



SO- 100- 



150 - 



60 - 200 - 



90 



120 



1501- 



250 



300 



350 - 



400 



450 



500^ 



Coal 



Coal 



Coal 




Waynesburg 
cool 



Coal 



LA. .li. .E- -UT^ 




Sewickley 
coal 



Pittsburgh 
cool 







ROD, 

pet 

40 60 80 100 

-\ r 



CORE 

RECOVERY, 

pet 

94 96 98 100 



Coal 



Coal 



Waynesburg 
coal 



Sewickley 
deool 



Pittsburgh 
eoal 



FRACTURE 

FREQUENCY, 

per meter 

2 4 6 



FIGURE 10. - Geotechnical logs for the Gateway site. 



strength of various strata in the over- 
burden. The geophysical data, in addi- 
tion, are useful for interpreting strati- 
graphic boundaries and changes O) . 

LABORATORY GEOPHYSICAL PROPERTIES 

Geophysical properties of coal measure 
strata, such as elastic wave velocities 
and electrical properties, are not well 
documented in the literature but are re- 
quired for the design of geophysical 
probes and interpretation of field geo- 
physical data. Hence, laboratory tests 
were conducted for these properties. 
Acoustic core logging was conducted on 
the core as it arrived from the field, 
before specimens were prepared for other 
laboratory tests. Measurements of P-wave 
traveltimes were made in different di- 
rections parallel to the bedding at 0.5- 
to 1-ft depth increments along the core. 



Velocity averages for the plane and ve- 
locity difference expressed in terms of 
anisotropy were then plotted as functions 
of depth and compared with the density 
and lithologic logs (fig. 12). As in 
field results, combinations of low veloc- 
ity and density normally signify coal 
seams and fractured zones . High values 
of velocity anisotropy usually suggest 
fracturing and poor quality of rock (4^) . 

Although electrical properties were de- 
termined only in the Gateway strata for 
the dry condition, hundreds of tests were 
conducted in various categories of coal 
measure rocks over frequency ranges from 
1 kHz to 100 MHz. Predicted propagation 
distance (skin depth) for electromagnetic 
energy over the frequency range from 1 
kHz to 20 MHz is indicated for the com- 
posite rock types in figure 13. 



23 




03 

■>. 

a 

a> 
a 
O 



I/) 
en 
O 



I/) 

Q. 
O 
0) 

O 



LU 

o 



24 



DENSITY, g*m' 



DIAMETRAL P-WAVE VELOCITY , km/s 



VELOCITY ANISOTROPY, pet 

20 30 00 50 




FIGURE 12. - Laboratory wave velocity and density profiles. 




5 I02- 



10' 102 10= 

FREQUENCY, kHz 



FIGURE 13, - Skin depth as a function of frequency, 

MECHANICAL PROPERTIES 

Mechanical properties were determined 
in uniaxial and triaxial compression 
tests using the closed-loop, servo- 
controlled testing machines and confin- 
ing pressure cell. Resultant data were 
plotted as axial, circumferential, and 



volumetric stress-strain curves (fig. 
14). Apparent Poisson's ratio, that is, 
the ratio of lateral to axial strain, was 
also determined throughout the range of 
stress. Concurrent with uniaxial com- 
pression testing, the change in P-wave 
velocity as a function of stress was 
determined (fig. 15). The effect of 
confinement on shear strength was deter- 
mined using conventional Mohr-Coulomb 
plots and analysis (fig. 16). Although 
stress-strain behavior in most of the 
coal measure rocks exhibited nonlinear 
behavior, the linear Coulomb failure cri- 
teria appeared to provide a good first 
approximation to shear strength for most 
of the rock types tested. Confining 
pressures ranged up to 2,500 psi (17.25 
MPa) , simulating expected lithostatic 
pressures to depths of roughly 2,750 ft. 
Typical triaxial test results are given 
in table 2. Enhancement of compressive 
strength with confinement is indicated in 
figure 17. More detailed analyses, pre- 
dicting failure by more complicated fail- 
ure criteria, are still underway. In- 
direct tensile strength also was deter- 
mined using the Brazilian test method. 





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CO 


3 3 l-l 
< 1-1 IM 






















CO 


4J 






















3 


rH 

3 


























4-1 M 






















CO 


03 


3 CO 






















4J 


0) 


<U 3 3 






















CO 


M 


•H iJ O 






















-3 




a (1) -H 


CN 


00 


o 


vi- 


o 


-<3- 


CO 


00 


CM 


t~~ 




J-l 


•H 4J 4J 


-* 


m 


in 


vo 


vO 


o- 


00 


r^ 


CJN 


1^ 


4J 


CO 


iM 3 O 


• 


• 


• 


• 


• 


B 


• 


• 


9 


• 


CO 


Q) 


14-1 -H -H 


o 




















(U 


4-1 


(U l-l 
O <4-l 4-1 






















4-1 


iH 


o o 






















(U 


CO 
•H 
























1-) 


> 






















X 


4-> 






















09 


CO 


3 3 






















3 


•H 


O (U 






















0) 


U 


•H -H 


u-1 


in 


o 


o 


o 


o 


in 


o 


o 


m 


4-1 


4-1 


CO O CO 


• 


• 


• 


• 


• 


• 


• 


• 


9 


• 






a) -H CM 


CM 


00 


vD 


r-^ 


r-~ 


<f 


00 


r-- 


r^ 


i-H 


3 


M-l 


x: 4-1 S 


f-H 




^H 


.— ) 


t — 1 


t-H 


t-H 


t-H 


t-H 


t-H 


CO 


O 


O 4-1 
O 0) 






















•H 
iH 


>> 


o 






















1-1 


u 


o 






















N 


























CO 
























^ 


4-1 






















U 


s 


O CO- 






















P3 


3 


(U "O 
























w 


l-l ^ 1) 






















T3 




(U a, 4-1 


(T. 


<-H 


CO 


1— 1 


<f 


r~ 


CO 


r^ 


00 


O 


3 


1 


^ d CO 


in 


f-H 


i-H 


CO 


CM 


-* 


t-H 


t-H 


in 


CO 


CO 




B CO 0) 
























• 


3 CO 4-1 






















M 


(M 


IS 






















3 
O 


w 


^ 






















ij 


CO 






















to 


^ 


CO 


r-. 


eo 


Cvl 


r^ 


CO 


-* 


t-H 


o 


CO 


vO 


CO 


H 


•O 3 4J 


9 


• 


• 


• 


• 




• 




• 


• 


p 




<U ^ 4-1 


■vT 


vO 


CN 


^ 


-0- 


r^ 


o 


CM 


vO 


t-H 


<x 




P9 CJ 


CNJ 




t-H 


CSl 


rH 


CM 


t-H 


t-H 


CM 


I-H 







•H 






















o 




X: 






















o 




4-1 






















l-l 






























o 


in 


<T\ 


in 


CM 


<3- 


00 


m 


r~~ 


v£> 


CO 




A 


• 


• 


■ 


• 


• 


• 


• 


• 


• 


• 


•H 




i-l 


o 


CJN 


CO 


CO 


CM 


vO 


00 


00 


vO 


in 


>4 




^ CO 


O) 


m 


m 


t^ 


<»■ 


t-H 


^ 


r^ 


CO 


o 


CO 




4-1 > 


f-H 


t-H 


CNj 


-* 




CM 


-* 


CO 


CO 


<r 


1-1 




O. H 4-1 


1 


1 


1 


1 


1 


1 


1 


1 


1 


1 


u 




0) 0) 4-1 


ro 


CO 


t-~ 


00 


ON 


O 


r~. 


in 


v£> 


o 


4-1 




n 4-1 


• 


• 


• 


• 


« 


• 


• 


• 


• 


• 






3 


in 


CO 


t-H 


00 


r-- 


ON 


CO 


NO 


o 


-* 


t3 




•H 


(J^ 


in 


o- 


<!• 


Csl 


00 


CO 


VO 


f—i 


ON 


3 








'~' 


CM 


-* 


t-H 


t-H 


-^ 


CO 


CO 


CO 


CO 


























i-l 


















• 




• 




CO 
















1 • 


• 




• 




iH 
















X) dJ 


• 




<u 




X 
















3 >-i 


0) 




3 




CO 




O 3 












CO to 


3 




O 




1-1 




•H O 












CO ^ 


o 




4J 




3 




bO -H 












CO 


4J 




CO 




3 




O 4-1 










<u 


TJ 


CO 




2 








.-( a 










iH 


(U 'V 


4J 




s 




CO 




O -H 


01 








CO 


TS 3 


tH 




1-1 


0) 


(U 




J3 M 


3 








,3 


T3 CO 


1-1 




tH 


3 


CO 




4J O 


o 








CO 


0) 


CO 






o 


•H 




•H CO 


4J 










^ 4) 






>N 


4J 


M 




I-] (U 


CO 








>. 


U 3 


>% 


<u 


0) 


CO 


& 




•a 


T3 


o 


O 


o 


T3 


0) o 


T) 


r-t 


rH 


01 









3 


Q 


Q 


a 


3 


4J 4J 


3 


CO 


CO 


e 


O 






CO 








CO 


3 CO 


CO 


J3 


Si 


1-1 


O 






en 








en 


M 


en 


CO 


en 


kJ 


~ 



25 



26 




□ Lateral strain 
A Volumetric strain 
V Axial strain 
O Apparent Poisson-s 
ratio 



4 6 



MILLISTRAIN 

FIGURE 14. - Volumetric stress-strain curve. 



0.2 



0.4 



0,6 



0.8 



PHYSICAL AND INDEX PROPERTIES 

Physical and index properties deter- 
mined in the laboratory included shore 
hardness, porosity, permeability, bulk 
density, point-load strength index, and 
slake durability. Moisture content was 
determined in the specimens tested 
for uniaxial and triaxial compression 
strength. In certain cases, studies also 
were made of the effect of moisture on 
elastic and strength properties and on 
the elastic wave velocities. 

STATISTICAL ANALYSES 

All rock property data were input into 
a computer for further analyses. Statis- 
tical analyses were conducted to obtain 
the mean, standard deviation, and coeffi- 
cient of variability for all measurements 
in a particular rock type at each site. 



Bar charts were constructed providing the 
mean and range (one standard deviation) 
of property values for each property 
(figs. 18-19). Because of lithologic 
variability within each rock category, 
standard deviations frequently were quite 
large. 

Crossplots were constructed to deter- 
mine correlations between rock properties 
by regression analyses for each rock type 
(fig. 20). Correlation coefficient ma- 
trices were obtained to indicate the 
degree of correlation between any two 
properties. Those properties that ex- 
hibit significant correlation with elas- 
tic moduli and strength properties and 
that could be easily and inexpensively 
obtained during exploration investiga- 
tions might, in some cases, be used as 
index properties to predict modulus or 
strength (fig. 21). 



27 



VELOCITY vs STRESS 



STRESS-STRAIN 



T 1 r 



Mudstone 




EL-7 No 303 



10 15 20 25 30 35 40 



40 
35 
30 
25 
20 
15 
10 
5 







I 


1 1 1 1 1 1 1 


p 


~ 


Mudstone ^ 


- 


- 


y Z\.-l No. 303 


- 


^-rr"- 


"1 1 1 1 1 1 1 





I 23456789 10 



"T — 1 — r 



1 — I — I — I — I — r 



Shale 




EL-7 No 125 B 



I I I I ] I J I J J 



10 20 30 40 50 60 70 80 90 100 110 



120 
100 
80 
60 
40 
20 





1 


1 1 1 


1 1 1 1 1 1 


r^- 


- 


Shale 


^^ 




^ 


"i 1 1 


EL-7 No. 125 B 
1 1 1 1 1 1 


- 



I 2 3 4 5 6 7 8 9 10 II 12 



4 - 



3 - 



1 
Coarse 


1 1 
-grain 


1 1 
sandstone 


— 


1 


' 1 


1 1 


EL-7 

1 1 


No. 


IIOA 

1 



80 
70 
60 
50 
40 
30 
20 
10 



1 1 1 1 1 1 1 1 1 


~~o- 


Coarse-groin sondstone X 


- 


j^ EL-7 No. IIOA 


_ 


^-°r^i 1 


" 



10 20 30 40 50 60 70 80 



23456789 10 




Shaley linnestone 



EL-7 No 296 



J I 1 I I I : L_ 



I8U 


1 


I 




1 1 I 1 


p 1 


160 


- 






y 


/ 


140 
120 


- 


Sh 


oley 


limestone /^ 


- 


100 


- 






y^ 


- 


80 


- 






y 


- 


60 
40 


- 






y EL-7 No, 296 


- 


20 


- 


-°^i 


^ 


1 1 1 1 


- 







1 




2 3 4 5 

MILLISTRAIN 


e 



20 40 60 80 100 120 140 160 180 
UNIAXIAL STRESS, MPa 

FIGURE 15. - P-wave velocity as a function of uniaxial stress. 



28 



-20 



20 



Micaceous 
sandstone 




40 60 80 100 120 140 

NORMAL STRESS, MPa 
FIGURE 16. - Mohr-Coulomb plot for the Pittsburgh Sandstone. 



180 200 220 



250 



o 

CL 


200 


2 




m 




CO 




UJ 




oc 




\- 


150 


U) 




_) 




< 




Q. 




O 




tr 


100 


Q. 




S 




Z3 




S 




V 




< 


bO 



-| I I I r 



Micaceous sandstone 



T I I F 



o 



^0* 



Principal stress equation, oj = III + 5.23 o-. 
Correlation coefficient, r = 0.96 



5 10 15 

CONFINING PRESSURE, MPa 
FIGURE 17. - Increase in compressive strength with confinement-Pittsburgh Sandstone. 



20 



29 



COMPRESSIVE STRENGTH. MPa 



BULK DENSITY, g/cm' 



Rock 


40 80 120 160 200 240 


Number 
tested 


Sandstone 


t 1 1 1 I 1 


21 


W////.k'<'////A 


Shole 




14 


1//////l^//////\ 


Siltstone 


V/Al/M 


5 


Mudstone 


■' V/I^M 


5 


Shaley limestone 




26 


[■////■//////♦■/////■/////J 


Silty shale 


ry^ 


4 


Limestone 




34 


l/////////^^/////////A 



Rock 


20 25 30 


Number 
tested 


Sandstone 


V/^/A 


105 


Sliole 


V/////l^////A 


64 


Siltstone 


* 


17 


Mudstone 


V/%'A 


IS 


Stioley limestone 


V/y^/A 


68 


Silty sltole 


Vj^ 


56 


Limestone 


V/^/A 


90 



INDIRECT I8RAZILIAN) TENSILE STRENGTH, MPa 



Rock 


1 3 5 7 9 II 


Number 
tested 


Sandstone 


■ 1 I 1 1 1 


65 


l'////////A/////////A 


Shale 


V//////^///M///////Af///////////////////M 


119 




Siltstone 


V/////%^////A 


3 


Mudstone 


V/W/M/IV////U/M 


33 


Shaley limestone 


w/////////l^////////A 


52 


Silty shole 


V////////J^///////M 


72 


Limestone 


U//////////^///////////A 


105 



POROSITY, pel 



Rock 


2 4 6 8 10 


Number 
tested 


Sondstone 


■ ■■III 


33 


v//////^//////x 


Shole 


f7///f////J 


39 


Siltstone 


V//^///A 


8 


Mudstone 


y///ifV/A 


8 


Shaley limestone 


V/////////^/////////M 


29 


Silty shale 


V///^///A 


38 


Limestone 


V/////////Af/////////A 


38 



STRAIN AT FAILURE, pet 



Rock 


Q2 04 06 08 10 1.2 


Number 
tested 


Sandstone 


1 1 I 1 1 1 


21 


W///W/W//Wk'///////W//////A 


Shale 




14 


V///////l^///////A 


Siltstone 




5 


y//////////^//////////i 


Mudstone 




5 


V////J^/////i 


Shaley limestone 




29 


v/////A/.'^///A 


Silty shale 




4 


VW///^/////M 


Limestone 




34 


V/////%/////,\ 



PERMEABILITY, dorcy 



Rock 


IO-« 10"' 10'* 


Number 
tested 


Sandstone 




27 






Shale 




19 






Siltstone 




4 


V//////////////////////^V///A 




Mudstone 


♦ 


1 


Shaley limestone 




13 






Silty shale 




14 






Limestone 




21 







STATIC YOUNG'S MODULUS, GPa 



Rock 


10 20 30 40 50 


Number 
tested 


Sandstone 




21 


V////////i%/W//////i 


Shale 




14 


V//i%///M 


Siltstone 


yJ^^ 


5 


Mudstone 


V/I^M 


5 


Shaley limestone 




29 


r///////l^'//WM 


Silty shole 


rz^T?, 


4 


Limestone 




34 


y//////4v/////A 



SHORE HARDNESS 



Rock 


10 20 30 40 50 60 


Number 
tested 


Sondstone 




64 


V//////AIM///A 


Shole 


V///////M'//////A 


115 


Siltstone 


\Zdk!A 


3 


Mudstone 


V////////J^>///////M 


32 


Shaley limestone 


Y////A////M 


53 


Silty shale 


l'////////A'////////A 


78 


Limestone 


V/////%////M 


108 



POISSON'S RATIO, stotic 



Rock 


05 10 015 20 25 30 


Number 
tested 


Sondstone 


r I 1 1 I 1 


17 


1//////<'^A(///<'WM 


Shale 




1 1 


V//////^//////M 


Siltstone 




4 


1////i%'////A 


Mudstone 




5 


'iy','///y/y//y//////y///y%//////////////^////^//» 




Shotey limestone 




27 


y////////^///////x 


Silty shole 




3 


y////^y///A 


Limestone 




32 


V/////^'/////i 



POINT LOAD INDEX, MPo 



Rock 


5 10 15 2 2.5 30 3 5 


Number 
tested 


Sondstone 




18 


W//////J^///////////X 




Shale 


* 


50 


Siltstone 




51 






Mudstone 




21 






Shaley limestone 




26 






Cool 


♦ 


Z 


Limestone 




19 


■///////////^///^//AAl/A////^A/,'//////y^////yM 





FIGURE 18. - Physical-mechanical property mean and standard deviation by rock type. 



30 



AXIAL P-.WAVE VELOCITY, km/s 



Rock 


1 2 3 4 5 6 


Number 
tested 


Sandstone 


w^y/A 


105 


Shale 


V///^///A 


64 


Silretone 


V/AfVA 


17 


Mudstone 


W///\////A 


18 


Shaley limestone 


Y/////!lt/////A 


65 


Silty sliole 


V///A,V//A 


56 


Limestone 


v//////Ay/////A 


83 



DYNAMIC YOUNG MODULUS, GPo 



Rock 


10 20 30 40 50 60 


Number 
tested 


Sondstone 




34 


f/'/V/V^y/V/y/i 


Shole 




24 


V///,\///M 


Stltstone 




6 


V///^'///A 


Mudstone 




7 


vm/z.^v/////! 


Sttdtey limestone 




27 


V//////////^//////////A 


Sllty sttole 




8 


V////^////A 


Limestone 




34 







AXIAL S-WAVE VELOCITY, km/s 



Rock 


0,5 1.0 15 20 25 3,0 


Numlier 
tested 


Sandstone 


1 1 1 ' ■ ■ 


39 


K////4y////i 


Shale 


W///A^/////X 


31 


Siltstone 


yy//////i(///////A 


5 


Mudstone 


y//////J^/////A 


10 


Stioley limestone 


V////>y///A 


22 


Sitty shole 


V/////Jlf'////A 


7 


Limestone 


W///My////A 


34 



DYNAMIC SHEAR MODULUS,GPo 



Rock 


4 8 12 16 20 24 


Number 
tested 


Sandstone 




34 


V/WZ.\/////A 


Shale 




24 


V////.^////A 


Siltstone 




6 


W///^////i 


Mudstone 




7 


V/////^'/////A 


Snoiey limestone 




27 


V/////////y///////^^A 


Silty shole 




8 


K/'///X^.f//////J 


Limestone 




34 


fy//'//'//'/V/fW/'/y/'/'///'l 



POISSON'S RATIO, dynamic 



Rock 


010 015 020 025 30 35 


Number 
tested 


Sandstone 




34 


y///////////^//MM///M 


Shale 




24 


'f////////ll////////A 


Siltstone 




6 


l'//////'////y///i^V/////////////A 


Mudstone 




7 


vw//////A^y//////////i 


Stioley limestone 




26 


ty/^///^'^/y'//i 


Silty stiole 




8 


Y/////////i^/////////A 


Limestone 




34 


V/////,^//////i 



FIGURE 19. - Acoustical property mean and standard deviation by rock type. 
ROCK CLASSIFICATION FOR ENGINEERING PURPOSES 



INTACT ROCK 

The strength and the modulus ratio 
were plotted according to the Deere (_5) 
intact-rock classification scheme that is 
widely used in civil engineering and 



mining (fig. 22). The scheme describes 
the rock in terms of five categories of 
compressive strength, and three catego- 
ries, high, medium, and low, of the ratio 
of modulus to strength. 



31 



•: 5 - 



en 



^"4 

o 

o 

_I 

UJ 

> 

> 3 

< 

5 





1 


1 1 
KEY 


1 


1 


1 
V 




V 


Limestone 


-sholey limestone 










▲ 


Mudstone- 


■siltstone 




V 






o 


Sandstone 




a 


2^^^.^.— - 


,^ — 




D 


Shale 










• 


Coal 


OV o ^^-^^ 


V V 


D 
V 






" 


D 
D C 


o v- — 


O 






~ 


▲ 


X 












a 


x^ 


▲ 










D 


/ 






y = x/ (0.3559+ O.I46x) 


- 




a 






r = 0.92 






/A 


D 

1 


1 1 


1 


1 


1 





2 3 4 5 6 

P-WAVE VELOCITY, INTACT ROCK,km/s 
FIGURE 20. - Crossplot of in situ versus laboratory P-wave velocity. 



AL 

RENGTH, MPa 

OJ X 

o c 
o c 


- 


1 


\ 1 1 


1 1 

D 

D 

D 
D 


^ 


UNIAX 
SSIVE ST 

O 
O 






a 






^ 100 


- 




1 1 1 




- 


Q. 

2 




a 


y= -42.0+45.2X 




O 




^^ 


r = 0.793 
1 i 









I 



6 



2 3 4 5 

P-WAVE VELOCITY, km/s 
FIGURE 21. - Crossplot of uniaxial compressive strength versus P-wave velocity. 



32 



Very low 
strength 



100 



o 

CL 
(D 

(n 

_l 
O 

o 



CO 

"o 

z 

o 



iLl 

o 

z 
< 



./*'* 






- / 

/ 
/ 



D 

Low 
strength 



.^^ 



/ <^^ 






/. 



/ 
/ 



>.° 



V 



Medium 
strength 



T — I I I 






/ 






. 8 



a o 

D 



.^o" 



J I I L 



B 

High 
strength 



Tf7 






"a^ 






Very high 
strength 



"I — TT 
/ 



"I — I I I 






v^ 



KEY 

o Sandstone 

D Shale 

V Limestone 

^ Shaley limestone 

A Mudstone 

▼ Silt stone 

■ Silty shale 



J I L 



10 



100 
COMPRESSIVE STRENGTH, MPa 

FIGURE 22. - Intact-rock engineering classification. 



1,000 



33 



Figure 22 shows that for the coal mea- 
sures at the Gateway Mine, the limestones 
and shaley limestones are classified 
in the high- to very-high-strength, and 
medium- to low-modulus-ratio categories. 
Siltstone and silty shales are predomi- 
nantly medium strength with low modulus 
ratio, and mudstone is low strength with 
medium to low modulus ratio. Such clas- 
sification schemes for intact rock are 
relevant to drilling, blasting, and 
fragmentation on a smaller scale, and 
for massive rock without joints. These 
schemes also aid in selection of appro- 
priate mining excavation and fragmenta- 
tion equipment. 

ROCK MASS 

Rock mass classification schemes take 
into account the influence of discontinu- 
ities and often, directly or indirectly, 
in situ environmental factors such as 
stress and moisture for estimating 
strength and def ormational behavior of 
rock masses. The geomechanics classifi- 
cation proposed by Bieniawski (6) appears 



adaptable to coal mining applications, 
and has been used to a limited extent in 
classifying roof conditions. The classi- 
fication is based on uniaxial compressive 
strength, RQD, the spacing, orientation 
and condition of joints, ground water 
conditions, and sometimes other modifying 
parameters. Significant parameters in 
the geomechanics classification for de- 
termining roof conditions were estimated 
for 50 ft of strata overlying the Pitts- 
burgh Coalbed. Roof strata at the Gate- 
way Mine were divided into three distinct 
lithologic units, and the rock mass rat- 
ing was applied to each lithologic member 
(table 3). The lowermost member was 
classified as poor rock, the middle mem- 
ber as good rock, and the uppermost mem- 
ber as fair rock. Such determinations 
permit speculation on maximum unsupported 
roof span and standup time and assist in 
selection of appropriate support. For 
application in coal measure roof rocks, 
however, modification and improvement of 
the classification scheme is needed to 
obtain a higher degree of predictability 
of standup time and support requirements. 



TABLE 3. - Rock mass classification of Gateway roof rock 



Lithologic member and thickness. . .ft. 



I, 10.8 



II, 24.7 



III, 19. 



Strength of intact rock: 

Uniaxial compressive strength. .MPa. . 

Rating 

Point-load index ' MPa . . 



73... 
(7).. 
0.08. 



RQD 

Rating. 



,pct, 



49., 
(8), 



106 , 

(12) , 

0.37-0.66.., 



98.. 
(20), 



Spacing of discontinuities m. 

Rating 



0.1-0.5. 
(10).... 



0.5. 
(20), 



Condition of discontinuities, 



Slickensided 



Rating. 



(6), 



Slightly rough, 

separation 

< 1 mm. 
(12) 



Ground water general conditions. 
Rating 



Moist. 
(7)... 



Moist. 
(7).., 



Total rating 

Class No 

Description 

Potential modifier; 
durability. 



slake 



pet. 



38 

IV 

Poor rock... 

75.5 (silt- 
stone) . 
91.1 (shale) 



71 

II 

Good rock. 



99.3. 



81. 

(7). 

0.09. 

97. 
(20). 

0.45-0.48. 
(20). 

Slightly rough, 

separation 

1-5 mm. 
(6). 

Moist. 
(7). 

60. 

III. 

Fair rock. 

87. 



•No rating is given because uniaxial compressive 
range. 



stength is preferred in the low 



34 



ROCK PROPERTIES DATA BASE 



Although the rock properties deter- 
mined in these investigations will be 
published, there is need for the es- 
tablishment of a computerized data base 
where rock property data can be obtained 
through search and retrieval opera- 
tions from remote terminal. The Bureau 
is working toward this goal by input- 
ting the property data into computerized 
files. The numerical data base manage- 
ment system, in addition, needs to be 
able to sort and retrieve coded numerical 



information so that the data can be 
manipulated and various mathematical and 
statistical analyses performed. As a 
start in this direction, property data 
tables from Bureau task files were orga- 
nized into a data base with standardized 
format for a wide variety of rock types 
(table 4). These property tables and a 
detailed description of the data base 
structure are being compiled into a Bu- 
reau Information Circular. 



TABLE 4. - Mechanical property tables for mine rock 



ID 


Rock type and 
modifier 


Location 

and 

description 


Source 


Young's 

modulus , 

GPa 


Pois- 
son's 
ratio 


Compressive 

strength, 

MPa 


Tensile 

strength, 

MPa 


Brazilian 

line-load 

strength, 

MPa 


1103 


Shale , 
calcareous 
kerogen. 


CO 


TCRC 


5.9 
(N=20) 




85.0 
(N=20) 


7.6 
(N=32) 










1104 


. . .do 


Garfield 
Co., CO. 


TCRC 


7.0 
(N=20) 


0.358 
(N=20) 


79.6 
(N=20) 






1105 


. . .do 


...do 


TCRC 


3.4 
(N=2) 


0.370 
(N=2) 


62.0 
(N=2) 






1106 


. . .do 


. . .do 


TCRC 


8.0 
(N=56) 


0.183 
(N=55) 


90.1 
(N=57) 


13.2 
(N=54) 




1107 


Shale. 


PA 


TCRC 


16.1 
(N=22) 




74.4 
(N=22) 




6.4 






(N=17) 


1108 


. . .do 


PA 


TCRC 


13.7 
(N=24) 




75.0 
(N=24) 




6.1 






(N=24) 


1109 


. . .do 


Rice Co. , 
KS. 


TCRC 


15.2 
(N=7) 




72.5 
(N=7) 






1110 


. . .do 


. . .do 


TCRC 


21.0 
(N=5) 




80.5 
(N=5) 







N 
TCRC 



Number of samples tested. 
Twin Cities Research Center. 



MINING APPLICATIONS 



35 



The physical and mechanical properties 
of coal measure and other mine rocks have 
application in most aspects of premine 
planning and mine design. Elastic and 
strength properties are especially needed 
in evaluating and modeling rock mass be- 
havior and structural stability in mines. 
Geologic data from both surface and un- 
derground surveys are required to deter- 
mine the continuity of coal seams and ore 
reserves, identify lithologic changes and 
trends, and delineate geological hazards 
in the proximity of mine workings. Geo- 
physical properties are necessary for 
interpreting rock mass and coal or ore 
characteristics in advance of mining and 
in inaccessible zones between exploration 
boreholes or adjacent to mine workings. 
Acoustic and electrical property data are 
particularly beneficial in the design of 
geophysical probes that identify and 
delineate rock mass conditions during the 
exploration phase of mining or in advance 
of the working face during mine develop- 
ment. Index properties are needed to 
infer other useful engineering proper- 
ties, when formalized testing procedures 



and facilities are unavailable or too 
complicated and costly for a mining op- 
eration to utilize on a cost-effective 
basis. Index properties can be particu- 
larly advantageous in coping with the 
frequent and common lithologic changes 
prevalent in coal measure rocks. More- 
over, the index properties determined in 
exploration or in-mine geotechnical pro- 
grams can be utilized in engineering 
classification schemes that infer rock 
mass response, indicate stability and 
support requirements , and permit experi- 
ence gained at one site to be transferred 
to other sites where similar conditions 
exist. 

Eventually, an interactive, computer- 
ized data base of engineering properties 
of rock can be established as an informa- 
tion retrieval system for use by mine 
operators, planners, and research or reg- 
ulatory agencies. Such a comprehensive 
engineering properties data base will 
require input from numerous sources and 
full cooperation of the industry. 



REFERENCES 



1. Lewis, W. E., and S. Tandanand 
(eds.). Bureau of Mines Test Procedures 
for Rocks. BuMines IC 8628, 1974, 
223 pp. 

2. Brown, E. T. (ed.). Rock Char- 
acterization Testing and Monitoring — 
ISRM Suggested Methods. Pergamon, 1981, 
211 pp. 



College Park, PA, June 11-14, 1972. Am. 
Soc. Civil Eng., 1972, pp. 649-687. 

5. Deere, D. V., and R. P. Miller. 
Engineering Classification and Index 
Properties for Intact Rock. U.S. Air 
Force Systems Command, Air Force Weapons 
Lab., Kirkland AFB, NM, Tech. Rep. AFWL- 
TR-65-116, 1966, 308 pp. 



3. Dresser Industries, Inc. Well Log 
Interpretation Techniques. 1982, 481 pp. 

4. Thill, R. E. Acoustic Methods for 
Monitoring Failure in Rock. Proc. 14th 
Symp. on Rock Mechanics, PA State Univ., 



6. Bieniawski, A. T. , F. Rafia, and 
D. A. Newman. Ground Control Investiga- 
tions for Assessment of Roof Conditions 
in Coal Mines. Proc. 21st Symp. on Rock 
Mechanics, Rolla, MO, May 28-30, 1980. 
Univ. MO— Rolla, 1980, pp. 691-700. 



36 



PILLAR DESIGN EQUATIONS FOR COAL EXTRACTION 
By Clarence 0. Babcock^ 



ABSTRACT 



Coal mine pillar design equations de- 
veloped during the period 1833 to 1980 
are reviewed. These equations suggest 
that mine pillars of different sizes are 
required for safety. Two widely used de- 
sign equations, those of Wilson and 
Wardell, give pillar areas that vary by 



an average factor of 2.08. The pillar 
width and height alone are not the pri- 
mary parameters in the design problem. 
The Mohn-Coulomb stress-failure criteria 
can be used to explain the difference in 
the estimated pillar sizes. 



INTRODUCTION 

Coulomb (J_)2 was the first person to 
publish a rational theory of earth pres- 
sures (1773), and this theory is in wide- 
spread use today for both soil and rock 
mechanics applications. His theory was 
also the first to show that the strength 
of a solid is related in part to the ma- 
terial properties and in part to the 
amount of constraint provided during the 
testing. Since that time, nearly every 



theory of failure of solids has been in 
terms of the combined stress, strain, or 
strain energy state. The role of con- 
straint in strength has only recently 
been emphasized in the design of mine 
pillars. Too often, the investigator has 
failed to realize that the constraint and 
not the material strength is responsible 
for pillar behavior when combined states 
of stress or strain are involved. 



STATE OF THE ART IN PILLAR DESIGN 



Wilson (2^) gave the equation for pil- 
lar size, W, in feet, for a safety factor 
(S.F.) of 1.0 as 

W = (R/3 + 2HD X 10"^) 

+ [R/3 + 2HD X 10-5)2 

+ (r2/3 - 4h2d2 X 10-6)]l/2^ (1) 

where R, H, and D are the entry width, 
entry height, and the depth of the coal 
seam below the surface, respectively, all 

in feet. 



the width and height of the pillar in 
feet, respectively. Wardell gives tables 
of minimum pillar widths for coal seams 
(pillar heights) of 4, 5, 6, 7, 8, 9, 10, 
and 12 ft, for a safety factor of 1.5. 
The pillar sizes are determined from a 
tributary area relationship that he does 
not define mathematically. A relation- 
ship that generates those tabulated val- 
ues was derived by Babcock and Hooker (4^) 
as 



(W + R)2 X 1.5 D 
W2 



1,000 _j_ 20 (W2) 



/H 



h2 



(3) 



Wardell (3^) proposed an equation of the 
form 



S = a/ »^ + b (W/H)2 

= 1,000 / /H + 20 (W/H)2 



(2) 



for the strength of actual mine pillars. 
Here the variable S is the strength of 
the pillar in pounds per square inch; a 
and b are coefficients; and W and H are 



where D is the depth below surface, R is 
the room width, and W and H are the width 
and height of the coal pillar, all in 
feet. 

^Mining engineer, Denver Research Cen- 
ter, Bureau of Mines, Denver, CO. 

^Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this paper. 



37 



Panek O) proposed an equation based on 
a theory of similitude for the design of 
mine pillars in which the physical prop- 
erties of the roof, floor, and coal are 



obtained from laboratory compressive 
testing of model pillars of the same 
geometry as the mine pillars with steel 
platens. His equation is 



Spred I cted _ ( ^r_ 
^ known \ ^s 



C4 



it 

Es 



C5 



Ucr 
Ucs 



where Spredlcted is the expected pillar 
strength, S^no^n is the measured model 
strength, and E is the Young's modulus, 
all in pounds per square inch; u is the 
coefficient of friction; and v is the 
Poisson's ratio. The subscripts c, f, r, 
and s denote the coal, floor, roof, and 
steel, respectively. The double sub- 
scripts cf , cr, and cs denote the co- 
efficient of friction between the coal 
and floor, coal and roof, and coal 



C6 



Ucf 
Ucs 



C7 



C8 



_1 



eg 



(4) 



and steel, respectively. The coeffi- 
cients C4 , C5 , ... C9 are constants to be 
determined by best fit statistical meth- 
ods. This relationship assumes no bond- 
ing between the coal and roof and floor 
rock. A more complicated relationship 
with three additional terms associated 
with C 1 , C2 , and C3 is given by Panek (_5) 
for the case when the model has a dif- 
ferent geometric shape than the mine pil- 
lar has . 



COMPARISON OF LABORATORY AND IN SITU SPECIMEN TESTING 
WITH OBSERVED MINE PILLAR BEHAVIOR 



In general, the pillar failure predic- 
tion equations are of the form 



^=A + B^ 



(5) 



where Op and Oq are the stresses at fail- 
ure in the pillar and the specimen; A, B, 
a, and 3 are constants; and W and H are 
width and height of the pillar, respec- 
tively. Numerous equations of this form 
and the results given in the literature 
were summarized by Babcock, Morgan, and 
Haramy (6^) and are shown in table 1 and 
figure 1. In figure 1, the zone between 
the failed and unfailed pillars was given 
by Warden O) for S.F. = 1.0. An empir- 
ical curve for S.F. = 1.5 is also shown. 
The theoretical result from reference 2 
(table 1) is so marked. 

It is apparent that the equations in 
the literature predicted as safe many 
pillars that were unsafe. While these 
equations may be correct for a given 
coal, none of them, other than Warden's, 
predicted correctly the large-scale or 
full-sized mine pillar behavior. 




123456789 10 
W/H RATIO 

FIGURE 1.- Comparison of pillar strengths pre- 
dicted by published equations and experimental 
strengths of mine pi liars and large test specimens. 
Source: Reference 6. (Underlined numbers in pa- 
rentheses refer to items in the list of references 
at the end of this paper. ) 



38 



33 



3 



+ 
< 



D 



B 
u 
o 

0) 



0) 

c 
o 
•w 

4-1 

nj 

3 

cr 

0) 

c 

M 
•H 
CO 
0) 

l-i 



I 






1 

)-l 


<u 


(1) 


o 


<w 


c 


(1) 


0) 


Pi 





r-» 00 ON o -^ — I 



M 
4J 

C 
3 

o 



Q) 

03 
CU 
U 

e 

0) 



<U 4J 



M 
O 
U 
CO 
bO 
1-1 
4J 
CO 
(U 
> 



u 
to 
(1) 






T3 

c 

C tS 

O M 



N <D 



Z CO 



0) <u 

c c 
o o 



CO CO O l-l 
0) 13 T3 X 

•H tO 



:^ 



bO 
3 
CO o 

u x) 



CO 
CO <U 

a w 
1-1 n) 
1-1 iJ 

:^^ 

U J-l 

3 1-1 
O C 



M 
l-i 
3 
Xl 

CO 
4J 
4J 
•H 

P- 

■V 

c 

(0 
^ • 

o 

(U 
PQ 



s 
o 

t3 
bO 
C 



CO 
O 

i-( 

x: o 

4-1 T3 

3 
O 

cn 



CO 

c u 

>. CO 
H ffi 

14-1 

>4-l >n 

3 0) 

O rH 

CO 

Q. C 
CD l-i 
(U (0 iH 

Q m pa 



CO 

3 
o 

■H C 

e CO o 

3 ^ T3 



B 
o 

T3 CO 
bO O 



o 
bO 
u 

CO 
►J 



< 0) 



CO TJ 



B 
o 

bO 



CO 
CO 

o 

z 



COC/)W3C/3cncrt|.Ji-JiJC/3C/lC/lCO(/lHJiJl-JHh4i-Ji-)CO 



IT) 

in u-1 CO 



vO u-1 



CM 

m lA CO 



-H sa- 



cs) ^^ ,-H O 



I I 



CM Ui 



o 



o o o u 



CJ 



o o 
in o 

l-s. CO 



00 -sT O O 

e» 00 o 
-* in vo 



o 

CSI 



PO CO 00 



o o o o 

u-1 O 



oooooooooo 



o o o o o 

^ o 
o 



^ 



• u 

• 0) 

• bO 

• 3 

• -l-l 

• XI 
4-1 O 
CO CO 

o 



0) 
C 4-> 

o c 

CO 0) 



O. 4-1 

M c 

CO 3 
> M ►-J U M 



3 XI 
CO O 



X rH 

4J (0 

•^ g 

<4-l 0) 

•H (I) 



• T) 

• 3 
>s CO 

-3 rH 

CO O 



crt o pd 



c 
o 
B 
to 

rH 

CO 
CO 



CO ^ r^ r-l ^ 

en r-s On O -H 
00 00 00 ^ On 



CM <Ti 

rH fO 
ON 0\ 



rH ^ VO <^ VO 

-* m in vo vo 

On On On On On 



• 3 

• 0) 

• T3 

• U 

• 0) 
3 0) 

O ts 



CO 



(U 
3 
3 bO 
iH CO CO 
S > S 



(M en -* 
r-- p^ r^ 

On ON ON 



T3 

iH 

<-{ 

3 
M 
4-) 
CO 

a 



in 

ON 



0) 
•3 
U 



ON 













o 














O 














T3 














(U 














U 














a 














CO 




• 










M 




B 










cu 




CO 










X 




a> 










4J 




CO 










o 




X 










rH 




o 










^ 




CO 










CO 




0) 










(U 




u 










rH 




o 










iH 




MH 










^ 




4J 














3 










«S 




(U 










X 




M 


• 




• 




4J 




0) 


0) 




Q) 




bO 




>4H 


,-\ 




rH 




3 


• 


IIH 


o. 


• 


a. 




a; 


vO 


•rA 


B 


0) 


B 




u 




"O 


CO 


rH 


CO 




u 


OJ 




en 


X 


en 




m 


CJ 


CO 




CO 








3 


iH 


u 


CJ 


4-1 




r^ 


<u 




CO 


•rl 


CO 




to 


u 


4-1 


0) 


rH 


(U 




CJ 


CD 


3 


4-1 


O. 4J 


• 


iH 


MH 


CO 




a 




>%x 


CU 


4-1 


0) 


CO 


rH 


u 


U 


Pi 


CO 


bO 




rH 


o 


3 




3 


V4 


4J 


CO 


a> 


CJ 




o 


CO 


o 


B 


X 




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o 


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z 


CO 


H 


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CO 

o 

o 


0) 
CJ 

u 

3 

o 


u 


t-I 


z 


CO 


H 


— 


CO 



39 



COMPARISON OF PILLAR SIZES DETERMINED WITH THE EQUATIONS 
OF WARDELL AND WILSON 



The equations of Wardell (2) and Wilson 
(1) were used to calculate the minimum 
pillar sizes required, in feet, for seam 
thicknesses of 6 and 10 ft, entry widths 
of 16, 20, and 24 ft, and depths of 200 
to 3,000 ft (table 2). The average pil- 
lar stresses corresponding to the pillar 
sizes in table 2 are given in table 3. 
Note that the stresses increase with 
depth. These results can all be repre- 
sented by straight line relationships as 
indicated by the parameters given in ta- 
ble 4. In this table, a, m, and r are 
the unconfined or uniaxial strength at 



the surface, the slope of the stress 
(ordinate) versus depth (abscissa) , and 
the correlation coefficient. If these 
results were a perfect fit to a straight 
line, r would have a value of 1.0. The 
values of 0.992 or better indicate that a 
straight line assumption is a reasonable 
one. Notice that by changing the in- 
tercept a and the slope, entry widths 
from 16 to 24 ft can be fitted. All 
these results could therefore be de- 
scribed by a Mohr-Coulomb model with rea- 
sonable success. 



TABLE 2. - Comparison of pillar widths, in feet, calculated 
with the equations of Wardell and Wilson 



Depth, 
ft 



16-ft room width 



Wardell Wilson 



20-ft room width 



Wardell | Wilson 



24-f t room width 
Wardell | Wilson 



6-ft-THICK COAL SEAM 



200 


21 


20 


23 


24 


26 


28 


400 


31 


23 


33 


27 


36 


31 


600 


39 


27 


41 


31 


44 


35 


800 


45 


30 


47 


34 


50 


38 


1,000 


50 


33 


53 


37 


55 


42 


1,200 


54 


36 


57 


41 


60 


45 


1,400 


59 


40 


61 


44 


64 


48 


1,600 


62 


43 


65 


47 


68 


51 


1,800 


66 


46 


69 


50 


72 


55 


2,000 


69 


49 


72 


53 


75 


58 


2,500 


75 


57 


80 


61 


83 


66 


3,000 


84 


64 


87 


69 


90 


73 







10-ft- 


THICK COAL SEAM 






200 


29 


22 


33 


26 


36 


30 


400 


46 


28 


49 


32 


52 


36 


600 


58 


33 


61 


37 


64 


42 


800 


68 


39 


71 


43 


74 


47 


1,000 


76 


44 


80 


48 


83 


52 


1,200 


84 


49 


87 


53 


90 


58 


1,400 


91 


54 


94 


59 


97 


63 


1,600 


97 


59 


100 


64 


103 


68 


1,800 


103 


64 


106 


69 


109 


73 


2,000 


108 


69 


112 


74 


115 


78 


2,500 


121 


81 


124 


86 


127 


91 


3,000 


132 


93 


135 


98 


139 


103 



40 



TABLE 3. - Comparison of average pillar stress, in pounds 
per square inch, calculated for the equations of Wardell 
and Wilson 



Depth, 
ft 



16-ft room width 
Wardell | Wilson 



20-ft room width 



Wardell | Wilson 



24-f t room width 



Wardell Wilson 



6-ft-THICK COAL SEAM 



200 


745 


778 


839 


807 


888 


705 


400 


1,103 


1,380 


1,238 


1,454 


1,333 


1,299 


600 


1,432 


1,826 


1,594 


1,949 


1,720 


1,778 


800 


1,764 


2,257 


1,951 


2,422 


2,103 


2,236 


1,000 


2,091 


2,646 


2,277 


2,848 


2,476 


2,615 


1,200 


2,420 


3,004 


2,628 


3,188 


2,822 


3,004 


1,400 


2,715 


3,293 


2,962 


3,554 


3,176 


3,372 


1,600 


3,039 


3,615 


3,283 


3,902 


3,514 


3,721 


1,800 


3,334 


3,924 


3,594 


4,234 


3,840 


4,016 


2,000 


3,642 


4,223 


3,919 


4,553 


4,182 


4,341 


2,500 


4,417 


4,921 


4,688 


5,290 


4,986 


5,094 


3,000 


5,102 


5,625 


5,445 


5,989 


5,776 


5,843 



10-ft-THICK COAL SEAM 



200 


578 


716 


619 


751 


667 


778 


400 


872 


1,185 


952 


1,268 


1,025 


1,333 


600 


1,172 


1,587 


1,270 


1,709 


1,361 


1,778 


800 


1,465 


1,909 


1,577 


2,061 


1,684 


2,191 


1,000 


1,758 


2,231 


1,875 


2,408 


1,994 


2,563 


1,200 


2,041 


2,534 


2,178 


2,732 


2,310 


2,878 


1,400 


2,323 


2,823 


2,471 


3,012 


2,614 


3,204 


1,600 


2,606 


3,103 


2,765 


3,308 


2,919 


3,514 


1,800 


2,883 


3,375 


3,052 


3,594 


3,216 


3,814 


2,000 


3,164 


3,642 


3,334 


3,873 


3,506 


4,104 


2,500 


3,846 


4,302 


4,046 


4,558 


4,241 


4,791 


3,000 


4,526 


4,945 


4,746 


5,219 


4,950 


5,473 



TABLE 


4. - Parameters for 


straight line fit to table 3 data 


Parameter 


16-ft room width 


20-ft room width 


24-ft room width 




Wardell | Wilson 


Wardell | Wilson 


Wardell | Wilson 



6-ft-THICK COAL SEAM 






505 


812 


615 


862 


684 


699 


m 


1.560 


1.682 


1.639 


1.809 


1.734 


1.796 


r 


.9995 


.9931 


.9993 


.9920 


.9989 


.9940 



10-ft-THICK COAL SEAM 



a 


329 


667 


387 


739 


445 


786 


m 


1.411 


1.476 


1.469 


1.553 


1.524 


1.636 


r 


.9999 


.9963 


.9997 


.9950 


.9995 


.9939 



The results given in tables 5 and 6 
show the need for horizontal confine- 
ment for pillar strength. The value d* 
is defined as d* = d/(l-re), where d is 
the depth below surface and r^ is the 
fractional recovery by mining (i.e., 
50 pet = 0.50). In table 5, the cohesive 
strength in unconfined shear is 100 psi. 
The small white area at the top of the 
table indicates that only small depths 
can be mined without some horizontal pil- 
lar confinement. For angles of internal 



friction of 30° or larger, a horizontal 
stress of less than 30 pet of the verti- 
cal is required for pillar stability away 
from the pillar edges. For large angles 
of friction, for instance, 50° or more, 
the horizontal stress required even for 
low 'a' values is only 11 pet or less of 
the vertical stress. This means that 
practically no horizontal confinement is 
required, and most of the confinement 
comes from the vertical stress loading 
across the failure surface. 



TABLE 5. - Blatio of horizontal pillar 
stress to vertical pillar stress needed 
for pillar stability for a small value 
of cohesive strength; 'a' = 100 psi 



d*, 




Friction 


angle (()>), 


deg 


ft 





10 


20 


30 


40 


50 


60 


200 


0.17 












^-^ = 
av 


400 


.58 


0.36 


0.20 


0.09 


0.03 




600 


.72 


.47 


.30 


.17 


.09 


0.03 




800 


.79 


.53 


.34 


.21 


.12 


.06 


0.02 


1,000 


.83 


.56 


.37 


.24 


.14 


.07 


.03 


1,200 


.86 


.59 


.39 


.25 


.15 


.08 


.03 


1,400 


.88 


.60 


.41 


.26 


.16 


.09 


.04 


1,600 


.90 


.62 


.42 


.27 


.17 


.10 


.04 


1,800 


.91 


.63 


.43 


.28 


.18 


.10 


.05 


2,000 


.92 


.64 


.43 


.29 


.18 


.10 


.05 


2,500 


.93 


.65 


.44 


.30 


.19 


.11 


.05 


3,000 


.94 


.66 


.45 


.30 


.19 


.11 


.06 


00 


1.0 


.70 


.49 


.33 


.22 


.13 


.07 



Next consider table 6 for a cohesive 
strength in shear of 800 psi. While this 
appears large, cohesion values up to 
1,240 psi for coal have been observed in 
the laboratory. Notice the large blank 
areas indicating that no horizontal con- 
finement is necessary. For angles of 
internal friction as small as 30°, a 
horizontal stress that is only 8 pet of 
the vertical is required for pillar 



41 



TABLE 6. - Ratio of horizontal pillar 
stress to vertical pillar stress needed 
for pillar stability for a large value 
of cohesive strength; 'a' = 800 psi 



d*. 


Friction angle ((Ji), deg 


ft 





10 


20 


30 


40 


50 


60 


200 
















400 
600 






^=0 
ov 










800 
















1,000 
















1,200 
















1,400 


0.05 














1,600 


.17 


0.01 












1,800 


.26 


.08 












2,000 


.33 


.15 


0.02 










2,500 


.47 


.26 


.12 


0.3 








3,000 


.56 


.33 


.18 


.08 


0.01 






00 


.00 


.70 


.49 


.33 


.22 


0.13 


0.07 



stability, away from the pillar edge 
where the opening stress concentration 
exists. For example, if a recovery of 50 
pet is used, the d* value would be 3,000 
for a depth of 1,500 ft. The <=° symbol at 
the bottom of tables 5 and 6 indicates 
the limits of horizontal constraint re- 
quired at great depth assuming elastic 
behavior still applies. 



THE ROLE OF CONSTRAINT IN PILLAR DESIGN 



WHAT IS FAILURE? 

What passes for coal or rock strength 
is often not material strength at all but 
constraint behavior. Consider the be- 
havior of the coalbed in figure 2 at some 
depth D below the surface. Regardless of 
the coal strength, the coalbed will re- 
main intact as long as it is confined by 
the rock strata above and below it. This 
will be true for any depth and any mate- 
rial. In other words, when completely 
confined, any coal or rock appears to 
have an 'infinite' strength. If a piece 
of this coal or rock could be tested in 
the laboratory, it would be readily ap- 
parent that the constraint and not the 
coal or rock strength was responsible for 
this behavior. 



Surface-N 



^^?<^^^<5^ 




^^^m 



Rock 

Vertical confining stress 

i i i J, i i J, i 1 i t i- 




FIGURE 2. - Confined coalbed. Any coal seam 
of any strength at any depth constrained as shown 
is "infinitely" strongfor a uniform vertical stress 



42 



Another way of showing the effect of 
constraint is to consider a single entry 
in the coal layer as shown in figure 3. 
If D is large enough and the coal in the 
rib is unconfined, it will break, adja- 
cent to the opening, for a distance that 
is some function of the coal seam height 
to point B. This broken zone will pro- 
gressively provide more constraint away 
from the rib until the combination of 
coal strength and constraint halts the 
breaking process. The constrained coal 
at A is unbroken. 

Next, consider several entries as shown 
in figure 4. Each entry will have two 
unconfined ribs in coal and an unconfined 
roof and floor. Because of less con- 
straint, a relatively greater volume of 

Surface^N, 



P^<^5^?^5^2s^J^?J^:?f^5^?^5^?^?;?:^J^?«?*5|^^ 



Rock 



Unconstrained at free 
surface of opening 



Constrained; 



rrrr 



Room 



JUJ, 



^Constrained 



A B C C B A 



^ 



t'T't^r t 



FIGURE 3. - Single entry in coal layer. Con- 
strained coal seam is "infinitely" strong except 
at the free surfaces and adjacent to them. 

Surface^ 



Rock 
Unconstrained at free surfaces 






Room kPiIIQ''^ Room SPillar? Room 



rfT 



A --Broken-' ^ 
^--Constrained-- 



FIGURE 4. - Several entries in coal layer. As 
the number of free surfaces increases, the overall 
constraint decreases, and the overall coal strength 
decreases as well. 



coal will break until the combined ef- 
fects of constraint and strength halt the 
breaking process. As the pillars become 
smaller, constraint plays a decreasing 
role in pillar stability and the coal 
strength without constraint a more impor- 
tant role. 

The many attempts to define pillar 
strength in terms of the width-to-height 
ratio (W/H) for the pillar imply that 
constraint is taken as necessary to en- 
sure pillar survival. If the coal alone 
is strong enough to support the applied 
load, constraint will be unnecessary and 
the value of W/H does not enter the prob- 
lem. What the W/H ratio represents is 
the amount and importance 
straint to pillar survival, 
en a physical meaning if 
Mohr-Coulomb behavior. 



of the con- 

This is giv- 

expressed in 



A common model for failure prediction 
is the Coulomb (J_) behavior of soil 
mechanics or the Mohr-Coulomb behavior of 
the theory of elasticity. Figure 5 shows 
the case of Coulomb failure in soil 
mechanics. The change in shearing 
strength with normal stress across the 
shear surface results from frictional 
effects, defined by the angle of internal 
friction ()>. That is, the shearing 
strength is the cohesive strength C in 
shear plus the frictional strength, a^ 
tan <t), that results from confinement of 
the shear surface by the normal stress 
On. If the normal stress is removed at 
any time, the cohesive strength in shear 



T5= C + CTp ton (^ 



Coulomb 
failure' 



1773 



Apparent 
strength 



Constraint 



Material 
I strengt h 



FIGURE 5. - Coulomb failure. Apparent strength 
as defined by the Coulomb a„ failure condition is 
partly material strength and partly the effects of 
constraint across the failure surface in shear. 



43 



is still C. In other words, the apparent 
increase in strength is only the effect 
of one stress, the shearing stress, act- 
ing against the confining stress. 

Another way of showing that the in- 
creased strength is not strength at all 
is shown in figure 6, which shows the 
frictional effect alone of one rigid sol- 
id surface being pressed against a second 
rigid solid surface. The resulting curve 
is the same as that for a Coulomb failure 
when the cohesive strength is zero. That 
is, the total apparent strength is only 
the result of constraint. Since the bod- 
ies are rigid they do not deform, and ma- 
terial strength is not considered. 

Next, consider the Mohr-Coulomb stress- 
failure relationship shown in figure 7. 
It is assumed in this theory that the 



C+cr, 



Coulomb failure 



Tg = CTn tan<^ , tan <^ = coefficient of friction 




A Shearing stresses produced when two rigid 
bodies are displaced parallel to the contact 
area. 




Constraint only, no 
material strength 



B The Coulomb relationship between the normal 
and shearing stresses oi A. 

FIGURE 6.. Shearing stresses and Coulomb fail- 
ure. The Coulomb failure relationship describes 
the effect of a shearing stress acting against a nor- 
mal stress across the shear zone. 




,-Mohr stress 
circle 



Apparent 
strength 

FIGURE 7. - Mohr-Coulomb stress-failure rela- 
tionship. This isof the same form as the relation- 
ship shown in figure 6 but with a cohesive shear 
strength added. 

Mohr's stress circle will produce failure 
if it touches the Coulomb failure line in 
any way. The Coulomb condition is a 
straight line as shown or a concave down- 
ward line when defined by the Mohr's 
stress envelope, with increasing dis- 
placement of the circle to the right on 
the normal stress axis. The strength of 
a solid is generally taken to be the val- 
ue of the shear stress at the point the 
circle touches the Coulomb surface. How- 
ever, in view of the fact that this is 
part shear strength and part confinement, 
this is an incorrect interpretation of 
pillar behavior. Since the solid can be 
broken with a shear stress C at any time 
when unconstrained, this much of the 
shear value is material strength and the 
remainder is constraint strength. The 
notion that the combined state of stress 
somehow measures strength has lead to 
confusion in the evaluation of pillar be- 
havior, particularly with regard to 
laboratory testing of rock and coal sam- 
ples, where unusually high constraint 
strength is provided by very stiff steel 
platens or triaxial confinement. 

The stress states at points A, B, and C 
in a coal seam as shown in figure 3 are 
defined in figure 8. If the coal has a 
shear strength C2, the Mohr's stress cir- 
cles A, B, and C will not reach the fail- 
ure condition, the rib will stand intact, 
and no constraint will be necessary for 
pillar stability. If the strength is C], 
the coal will break at the rib until the 
broken coal provides enough constraint or 



44 




FIGURE 8. - The Mohr's stress circles for 
points A, B, and C shown in figure 3. The coal 
with shear strength C, has failed at the rib, and 
constraint is necessary to pillar survival. The 
coal v/ith shear strength C2 does not fail at the 
rib, and constraint is unnecessary to pillar survival. 

the stress concentration disappears so 
that the Mohr's stress circle B will just 
touch the Coulomb failure condition. 
Constraint will be necessary for pillar 
survival. 

DOES A LARGE W/H RATIO ALWAYS ENSURE 
PILLAR STRENGTH? 

At least one implicit assumption inher- 
ent in the use of the W/H ratio as a mea- 
sure of pillar stability is that this ra- 
tio provides constraint to most of the 
pillar volume away from the coal rib. In 
part, this belief is the result of test- 
ing flat rock or coal samples in the 
laboratory between steel platens that 
provide frictional constraint to the sam- 
ple, making it almost incompressible. 

Consider the relationships shown in 
figure 9 where steel cubes load a rock 
or coal cube. For 30 psi of vertical 
stress in the steel, a vertical strain 
of 1 uln/in results. If the Poisson's 
ratio is 0.25, the horizontal strain is 
0.25 pin/in. This is an upper bound for 
the horizontal strain because the steel 
block is not constrained horizontally as 
would be the case for a steel platen that 
is much larger than the specimen. The 
cube of coal with a Young's modulus of 




Strong 



= Weak 




I^t MCTy 



Coal is 
constrained 
by shear 
stresses 




Strong 



FIGURE 9.- Steel cubes loading a coal cube. 
If the coal is confined between two more rigid ma- 
terials such as steel platens or strong roof and 
floor rock, the apparent strength of the coal in- 
creases from constraint. 

1 X 10^ psi or less would have a vertical 
strain of 30 yin/in and a horizontal 
strain of 7.5 yin/in for a Poisson's ra- 
tio of 0.25. The strains in the coal 
would be 30 times those in the steel. 
The net effect is that the steel con- 
strains the horizontal expansion of the 
coal and gives it an apparent strength 
that is much greater than the strength of 
the coal when unconfined. 

Next consider the relationships shown 
in figure 10 where cubes of rock with low 
Young's moduli, for example, 0.1 x 10^ 
psi, are used to load the cube of coal 
with Young's modulus of 1 x 10^ psi or 
less. A vertical stress of 30 psi will 
produce a vertical strain of 300 uin/in 
in the rock cubes. If the soft roof and 
floor are a claylike material, the Pois- 
son's ratio may be very high, approaching 
0.5. To be conservative, let the Pois- 
son's ratio be 0.25. This would result 
in a horizontal strain in the roof and 
floor of 75 pin/in. This is an upper 
bound because the roof and floor are con- 
strained laterally in the mine. For this 



45 




Weak- 




Flow 



U t t { ^y 



Strong 



Weak- 




Coal is not 
constrained 



Flow 



FIGURE 10. - Strong coal between weaker ma- 
terials. If the coal is stronger than the roof and 
floor rock or the material used to load the coal in 
the laboratory, the coal is not constrained. In the 
mine the roof and floor must flow for the constraint 
to disappear. 



horizontal strain to occur, the floor 
must heave and the roof must sag. These 
conditions are not uncommon in actual 
mining operations. The vertical strain 
in the coal would be 30 uin/in. The hor- 
izontal strain in the coal for a Pois- 
son's ratio of 0.25 would be 7.5 pin/in. 
The horizontal strains in the roof and 
floor would be 10 times that for the 
coal. The net effect is that, instead of 
confining the coal seam, the roof and 
floor would actually pull the coal seam 
apart. In this case the W/H ratio does 
not indicate increased strength. 

Future work on the effects of W/H ratio 
on pillar strength should evaluate these 
two important factors in terms of Mohr- 
Coulomb behavior. First, the strength of 
the coal itself, when unconfined, must be 
determined; and second, the role of con- 
straint or lack of it must be evaluated. 
One way of doing this, in situ, would be 
to use the drilling yield method devel- 
oped in Europe. 



POTENTIAL FOR USING THE DRILLING YIELD METHOD FOR DESIGNING MINE PILLARS 



Kidybinski and Stranz (26) reported on 
the drilling yield method to the U.S. De- 
partment of the Interior. In this method 
a hole drilled into the coal seam is used 
to study the conditions of stress in the 
seam. The volume of drill cuttings is 
expected to be 2 to 3 L/m of drill-hole 
length when the hole diameter is 42 mm. 
If the hole squeezes shut during drill- 
ing, indicating that the coal is not 
strong enough to handle the stress, the 
volume of cuttings will increase. If the 
volume is 6 L or more, a serious coal 
bump condition exists. The method has 
been in field use since the 1960's ( 27 ) 
in the Federal Republic of Germany, since 
the 1950's in the U.S.S.R. , since 1965 in 
Poland, and since 1968 in Yugoslavia. A 
2.5-hp motor was used, and the stalling 
behavior of the motor was also a part of 
the analysis. 

The use of the drilling yield method 
could be helpful in the design of 
mine pillars by establishing the broken 
or squeezing zone locations or depths. 
The concept is shown in figure 11. In 



part A, a point in the coal seam de- 
noted by P is confined and stable. If a 
hole is drilled through this location 
(part B ) , the constraint is removed on 
the hole surface. The tangential stress 

Roof 




Floor 
A Confined coal seam is stable at point P. 



Roof 



Drill 
hole- 




Floor 
B Unconfined coal seam may be unstable 
at point P. 

FIGURE 11. - Use of drilling yield method to 
establish pillar dependence on constraint for 
survival. 



46 



concentration on the hole boundary would 
be in the range of 2 to 3 times the ver- 
tical stress for an elastic condition. 
If the coal breaks around the hole, it 
does not mean that the pillar is unstable 
but that the coal strength is less than 2 
to 3 times the most compressive prin- 
cipal stress. With breaking, the stress 
concentration will be reduced and in the 



limit approach the coal seam stress. 
The breaking should then stop. If it 
does not , this indicates that the coal 
strength is inadequate to support the 
stresses when unconfined. The volume of 
cuttings is therefore an indicator of the 
pillar stability versus depth from the 
rib. 



SUMMARY AND CONCLUSIONS 



The number of studies of pillar behav- 
ior based on testing of samples has in- 
creased rapidly during the last 25 yr. 
The concepts on which many, if not most, 
of these tests were based often date 
back to 1912 or before. That is, the 
number of testing procedures or methods 
of analysis used has not kept pace with 
the number of studies. Recent experimen- 
tal testing has emphasized the effects of 
sample size and conditions of testing, 
including constraint. One reason for the 
slow progress in pillar design is that 
the separate effects of coal strength 
and the apparent increase in strength re- 
sulting from confinement have not been 
appreciated. 

In pillar design, if the pillar must 
have constraint from the roof and floor 
to survive, the pillar becomes unstable 
if such constraint disappears with time 
or change in mine geometry. If the coal 
can survive in pillar form without con- 
straint, it is more likely to remain sta- 
ble when constraint conditions are 
changed. 

The practice of assuming that a flat 
coal pillar will be "infinitely" strong 
because this is the case for laboratory 
testing of coal between steel platens 
should be examined carefully. In such 
tests, the strength of the platens and 
not of the coal is determined, and there 
is no correlation to mining conditions 
with roof and floor of rock, sometimes 
rock that is not very strong or struc- 
turally stable. 



It is commonly considered that the re- 
cent emphasis on constraint is new, but 
in fact, any equation using W/H relation- 
ships implies this condition. 

Most of the pillar strength results 
from constraint across the failure sur- 
face by normal compressive stresses 
according to the Mohr-Coulomb stress- 
failure criteria. This constraint is 
supplied for the most part by the ver- 
tical component of stress from the over- 
lying rock. This is particularly true if 
the angle of internal friction is 30° or 
more. For angles as large as 50° or 60° 
only a very small horizontal stress com- 
ponent is needed for pillar stability. 
In addition, the small value of the un- 
confined shear strength often assumed for 
in situ behavior is largely responsible 
for the very different results obtained 
in the laboratory and in the mine. 

Nearly all the theories with width- 
to-height relationships can be represent- 
ed equally well using the Mohr-Coulomb 
stress-failure models. There is no magic 
in the width-to-height ratios for pillar 
design. These occur naturally when the 
strength increases with depth and a given 
entry size is used. 

The cores of pillars are confined for 
the most part not by the horizontal pil- 
lar stresses, as in the constrained core 
concept, but through the action of the 
vertical stress away from the pillar 
edges. The effect that exists with 
respect to horizontal stress is one of a 
confined edge rather than a core. 



47 



REFERENCES 



1. Coulomb, C. A. Essai sur une Ap- 
plication des Regies des Maximis et 
Minimis a Quelques Problems de Statlque 
Relatifs a 1 'Architecture (Tests for the 
Application of Maxima and Minima Rules 
and Statistical Problems Relative to Ar- 
chitecture). Mem. Acad. R. Pres. Divers 
Savants (Paris), v. 7, 1773. 

2. Wilson, A. H. , and D. P. Ashwin. 
Research Into the Determination of Pillar 
Size. Part I. An Hypothesis Concerning 
Pillar Stability. Min. Eng. (N.Y.), v. 
131, June 1972, p. 409-417. 

3. Warden, K. , and Partners (Newcas- 
tle, United Kingdom). Guidelines for 
Mining Near Surface Waters (contract 
H0252021). BuMines OFR 30-77, 1977, 59 
pp.; NTIS PB 264 729. 

4. Babcock, C. C, and V. E. Hooker. 
Results of Research To Develop Guidelines 
for Mining Near Surface and Underground 
Bodies of Water. BuMines IC 8741, 1977, 
17 pp. 

5. Panek, L. Estimating Mine Pillar 
Strength From Compression Tests. Trans. 
Soc. Min. Eng. AIME, v. 268, 1980, 
pp. 1749-1761. 

6. Babcock, C. 0., T. Morgan, and K. 
Haramy. Review of Pillar Design Equa- 
tions Including the Effects of Con- 
straint. Paper in 1st Annu. Conf. on 
Ground Control in Mining, WV Univ. , Mor- 
gantown, WV, July 1981. Dep. Min. Eng., 
WV Univ., 1981, pp. 23-34. 



9. Johnson. Materials of Construc- 
tion. 1897. (Cited in Gonnerman, H. F. 
Effect of Size and Shape of Test Speci- 
mens on Compressive Strength of Concrete. 
Structural Materials Res. Lab., Bull. 16, 
Oct. 1925, 18 pp.; ASTM, v. 25, pt. 2, 
1925.) 

10. Carpenter, R. C. Article in Sib- 
ley J. Eng., V. 16, No. 3, Dec. 1901, 
p. 105. (Cited in reference 11.) 

11. Bunting, D. Chamber-Pillars in 
Deep Anthracite-Mines. Trans. AIME, v. 
42, 1912, pp. 236-245. 

12. Griffith, W. , and E. T. Conner. 
Mining Conditions Under the City of 
Scranton, Pa. BuMines B 25, 1912, 89 pp. 

13. Greenwald, H. P., H. C. Howarth, 
and I. Hartmann. Experiments on Strength 
of Small Pillars of Coal in the Pitts- 
burgh Bed. BuMines TP 605, 1939, 22 pp. 

14. . Experiments on Strength of 

Small Pillars of Coal in the Pittsburgh 
Bed. BuMines RI 3575, 1941, 7 pp. 

15. Steart, F. A. Strength and Sta- 
bility of Pillars in Coal Mines. Chem. , 
Metall. , and Min. Soc. S. Afr. , v. 54, 
1954, pp. 307-325. 

16. Gaddy, F. L. A Study of the Ulti- 
mate Strength of Coal as Related to the 
Absolute Size of Cubical Specimens Test- 
ed. Bull. VA Polytech. Inst, and State 
Univ., No. 49, 1954, pp. 1-27. 



7. Vicat, L. J. Researches on Physi- 
cal Phenomena Which Precede and Accompany 
Rupture or Deformation of a Certain Class 
of Solids. Ann. Fonts et Chaussees, pt. 
2, 1833, p. 201. 



17. Holland, C. T. The Strength of 
Coal in Mine Pillars. Paper in Proc. 6th 
Symp. on Rock Mechanics, Univ. MO, Rolla, 
MO, Apr. 1964. Univ. MO~Rolla, 1964, 
pp. 450-456. 



8. Bauschinger, J. Mitteilungen aus 
dem Mechanisch-Technischen Laboratorium 
der K. Technischen Hochschule in Munchen 
(Reports From the Mechanical-Technical 
Laboratories of the K. Technical College 
in Munich). V. 6, 1876. 



18. Evans, I., and C. D. Pomeroy. The 
Strength, Fracture and Workability of 
Coal. Pergamon, 1966, 277 pp. 



48 



19. Salamon, M. D. G. , and A. H. 
Munro. A Study of the Strength of Coal 
Pillars. J. S. Afr. Inst. Min. and 
Metall,, V. 68-2, Sept. 1967, pp. 55-67. 

20. Bieniawski, Z. T. In Situ 
Strength and Deformation Characteristics 
of Coal. Eng. Geol. (Amsterdam), v. 2, 
1968, pp. 325-340. 



21. 



In Situ Large Scale Test- 
Paper in Proc. Conf. on In 
and Rock. 



ing of Coal. 

Situ Investigations on Solids 

Brit. Geotech. Soc. , 1969, pp. 67-74. 



22. Van Heerden. In Situ Determina- 
tion of Complete Stress-Strain Character- 
istics for 1.4 M Square Coal Specimens 
With Height to Width Ratios of Up to 3.4. 
Rep. CSIR (S. Afr.), No. ME 1265, 1974, 
p. 30. 

23. Wagner, H. Determination of Com- 
plete Load Deformation Characteristics 
of Coal Pillars. Paper in Proc. 3d 
Int. Conf. on Rock Mechanics, Denver, 
CO. Natl. Acad. Sci., v. 11-B, 1974, 
pp. 1076-1082. 



24. Hustrulid, W. A. A Review of Coal 
Pillar Strength Formulas. Rock Mech. , 
V. 8, 1976, pp. 115-145. 

25. Skelly, W. A., J. Wolgamott, and 
F. Wang. Coal Mine Pillar Strength and 
Deformation Prediction Through Laboratory 
Sample Testing. Paper in Proc. 18th 
Sjnmp. on Rock Mechanics, Keystone, CO, 
June 1977. CO School Mines Press, 1977, 
pp. 2B5-1 to 2B5-5. 

26. Kidybinski , A., and B. Stranz. 
Coal Mine Safety Hazards Related to Rock 
Stresses. ' Res. Prog. Rep. — P. L. 480 to 
U.S. Dep, Interior by Polish Central 
Inst, of Mining (Katowice, Poland), Dec. 
1972, 56 pp. 

27. Jahn, H. (Identification and Dis- 
posal of Dangerous Stresses in the Coal- 
side of a Gateroad in the Coal Bump 
Prone Seam Sonnenschein. ) Gluckauf, 
1965, p. 101. 



49 



UNDERHAND CUT-AND-FILL STOPING FOR ROCK BURST CONTROL 
By F. Michael Jenkins^ and K. Robert Dormant 



ABSTRACT 



The occurrence of rock, bursts in deep 
metal and nonmetal mines presents a major 
hazard to their safe and economical oper- 
ation. As part of a Bureau of Mines pro- 
gram to reduce rock bursts in deep vein 
mining, a project was conducted in the 
Coeur d'Alene mining district of Idaho to 
evaluate and demonstrate the mining of a 
destressed sill pillar using underhand 
cut-and-fill methods. A 50-ft sill pil- 
lar in a burst-prone area was precondi- 
tioned by drilling and blasting vertical 



holes in the ore. After raises were 
driven through the destressed pillar to 
the level above, it was mined by the un- 
derhand cut-and-fill method. The con- 
clusion from this demonstration is that 
the combination of ore preconditioning 
and underhand mining resulted in great- 
ly improved rock burst and ground con- 
trol, allowing safe and efficient mining 
of a potentially hazardous, burst-prone 
pillar. 



INTRODUCTION 



The occurrence of rock bursts in deep 
metal and nonmetal mines presents a major 
hazard to their safe and economical oper- 
ation. A rock burst is the sudden vio- 
lent release of strain energy stored in 
the rock as a result of mining. The 
causes of rock bursting can be traced to 
the geometry of mine openings and rock 
characteristics. Strong, brittle rock in 
high stressed areas of the mine has a 
high potential for bursting. Bursting 
associated with stoping usually occurs in 
sill pillars, converging stopes, initial 
or subdrifts, and raises. 

As part of a research program to re- 
duce rock bursts in deep vein mining, a 



project was conducted under a Bureau con- 
tract to evaluate and demonstrate the 
mining of a destressed sill pillar using 
underhand cut-and-fill methods. 3 The 
project was carried out in two phases: 
first, to study the feasibility and cost 
effectiveness of the method and present 
design recommendations, and second, to 
enlist the cooperation of a mine and dem- 
onstrate that the method could be used to 
reduce rock bursts. This report presents 
some background information from phase I 
and the findings and conclusions of phase 
II as demonstrated by the test stopes, 
including cost, production, and safety 
studies. 



BACKGROUND 



In the Coeur d'Alene mining district of 
northern Idaho, where lead-zinc silver 
veins are being mined to 8,000 ft below 
the surface, rock bursting continues to 
be a severe operational problem. Finding 
a means of controlling rock bursting has 
been the goal of research for almost 80 
yr. However, the problem still persists 

^Mining engineer. 
^Supervisory mining engineer. 
Spokane Research Center, Bureau of 
Mines, Spokane, WA. 



because of its complexity and the diffi- 
culty of finding practical solutions that 
can be applied economically. 

The predominant mining method in the 
Coeur d'Alene mining district is overhand 

^Bush, D. D., W. Blake, and M. P. 
Board. Evaluation and Demonstration of 
Underhand Stoping To Control Rock Bursts. 
BuMines contract H0292013; for informa- 
tion, contact F. M. Jenkins, TPO, Spokane 
Research Center. 



50 



cut and fill. The general plan is to de- 
velop horizontal levels from vertical 
shafts, produce ore from stopes that are 
mined upwards toward the mined-out level 
above, and backfill each stope cut with 
mill tailings. The result is an ever- 
decreasing pillar of ore between the lev- 
el above and the miners working in the 
stope, with lateral stress concentrated 
in the pillar. Structural failure of the 
sill pillar often produces rock bursts. 

Experience indicates that sill-pillar 
bursting usually begins when the pillar 
height is reduced to about 80 ft, with 
the greatest frequency and severity oc- 
curring at pillar heights of 40 to 30 ft. 
These bursts often inflict great damage 
to the stope as well as to haulage drifts 
and crosscuts above and below. A large 
burst can displace more than 1,000 tons 
of rock, heave the drift floor, and break 
numerous caps and posts. The results 
are extensive repair costs and loss 



of production. Statistics indicate that 
failure associated with sill pillars ac- 
counts for more than 60 pet of all rock 
bursts.^ 

Ore preconditioning, the destress 
blasting of a pillar prior to mining, has 
been shown to effectively reduce rock 
bursts. 5 Blasting changes the character- 
istics of the ore from a strong, brittle 
material to a yielding material incapable 
of storing strain energy. To be effec- 
tive, however, the preconditioning must 
thoroughly fragment the ore. This cre- 
ates a hazard to those working in over- 
hand stopes because they are working be- 
neath fragmented rock. The underhand 
cut-and-fill mining method was developed 
in Canada to mine pillars of ore similar 
to those created by preconditioning. ^ A 
combination of the two techniques (pre- 
conditioning and underhand cut and fill) 
seemed an attractive solution to recover- 
ing burst-prone sill pillars. 



PHASE I— EXAMINATION AND DESIGN RECOMMENDATIONS 



During phase I of this project, the un- 
derhand cut-and-fill practice was ex- 
amined at two mining operations in Canada 
and two in the United States where the 
method was being used for primary pro- 
duction and for pillar recovery under 
difficult ground conditions. An artist's 
conception of a typical underhand cut- 
and-fill stope is shown in figure 1. 

Stress analyses were made of mining 
by overhand cut and fill and by under- 
hand cut and fill to compare their ef- 
fects on ground control in the Coeur 
d'Alene District. 

To examine the bursting potential of 
underhand cut-and-fill stoping, the ini- 
tial simulation assumes five end-to-end 
stopes mined in a flat-back arrangement 
toward an unmined level below. This sim- 
ulation ignores active mining on multiple 
levels; thus, no sill pillar is created. 
The maximum stress (30,000 to 50,000 psi) 
occurs at the stoping horizon. With no 
sill pillar being created, the stress 
does not change as mining progresses. 
Because of the constant rate of energy 



release, little bursting should be en- 
countered with a flat-back underhand 
method if no pillars are created. 

Because mining is not normally confined 
to one level, more realistic, multilevel 
mining sequences were examined. A series 
of simulations with simultaneous mining 
on three levels was made. The potential 
for bursting approaches that of the over- 
hand method soon after multilevel mining 
occurs. In this case, the pillar stress 
is between 55,000 and 60,000 psi or near- 
ly that amount induced in the sill pillar 

^McLaughlin, W. C, G. C. Waddell, 
and J. C. McCaslin. Seismic Equipment 
Used in Rock Burst Control in the Coeur 
d'Alene Mining District, Idaho. BioMines 
RI 8138, 1976, 27 pp. 

^Karwoski, W. J., W. C. McLaughlin, 
and W. Blake. Rock Preconditioning To 
Prevent Rock Bursts — Report on a Field 
Demonstration. BuMines RI 8381, 1979, 
45 pp. 

^Society of Mining Engineers of AIME. 
Underground Mining Methods Handbook. New 
York, 1982, p. 631 . 



51 




-L 




^9^ 




<*'>' iw'TSi'- 



FIGURE 1. - Artist's conception of an underhand cut=and-fill stope. 



for the overhand case. Though energy re- 
lease rates are lower, bursting will not 
be eliminated by underhand mining alone. 
However, underhand cut and fill offers 
distinct advantages where destressing is 
used in rock burst control: (1) Precon- 
ditioning techniques are easily applied 
since destress holes can be drilled ver- 
tically in the ore body, (2) damage dur- 
ing preconditioning will generally be 
less severe and will not produce the mas- 
sive caving that often occurs during con- 
ventional destress blasting, and (3) 
blast-induced caving and timber damage 
during mining are eliminated. 

Following the mine examinations and 
stress analysis studies, three alternate 



methods of underhand cut-and-fill stoping 
were suggested for the Coeur d'Alene. A 
cost-and-production estimate was made for 
the three methods as well as for typical 
timbered cut and fill. They indicated 
that efficiencies (tons per worker-shift) 
for the underhand cut-and-fill method 
were comparable to practices in use, and 
that considerable cost savings could be 
obtained over conventional timbered cut 
and fill, owing to the reduction of tim- 
ber required for support. The cost esti- 
mate found that underhand cut and fill 
would not, however, be an economical al- 
ternative for primary production at all 
mines. 



52 



PHASE II~FIELD TEST 



The phase I study indicated underhand 
cut-and-fill stoping could be a cost- 
effective method of rock burst control 
when used in conjunction with ore pre- 
conditioning. During phase 11, the meth- 
od was tested in an operating mine, and 
productivity, relative cost, and ground 
control were evaluated. This section 
discusses the field test, from site se- 
lection through mining of the final cut. 

SITE SELECTION AND MINING PLAN 

The results of the phase I study were 
presented to the major mining companies 
in the Coeur d'Alene District. The re- 
sponse was generally positive, although 
most of the companies questioned the eco- 
nomics of the system. One company was 
interested enough to cooperate with the 
Bureau and Terra Tek, Inc., in a field 
test. An ideal test area was selected in 
an unusual manner. Miners were preparing 
a relatively stable area for the test 
when a major burst occurred in a 50-ft 
sill pillar in another part of the mine. 
Based on past experience, management pre- 
dicted that additional pillar bursts of 
possibly greater magnitude were likely to 
occur during mining of this sill pillar. 
They chose this site for testing the un- 
derhand cut and fill combined with ore 
preconditioning. 

A mining plan, illustrated in figure 2, 
was prepared by Terra Tek and submitted 
to the Bureau and the mine for approval. 
The plan called for the following se- 
quence of events: 

1. Complete repair of the rock burst 
and complete the present stope cut. 
Three of the stopes in the pillar were to 
be brought to the same elevation, thus 
creating a flat back. One stope had been 
dropped because of poor-grade ore. 

2. Clean out and repair the haulage 
lateral above the pillar. Timbersets 
were to be repaired or replaced, slabs 
were to be removed, and track was to be 
repaired where possible. 



3. Prepare the haulage drift for ce- 
mented fill after cleanup and repair. A 
floor mat, consisting of 12- by 12-in 
stringers, lagging, 4- by 4-in No. 8 wire 
mesh, and a layer of woven polyethylene 
cloth (Fabrene),^ would be laid over the 
entire 600-ft length of drift. At 100-ft 
intervals, fill fences would be con- 
structed from light timber, wire mesh, 
and Fabrene to limit the extent of any 
individual pour. 

4. Drill the pillar with vertical de- 
stress holes and blast using ammonium 
nitrate-fuel oil (ANFO) or water gel. 

5. Perform pre- and post-precondi- 
tioning seismic velocity surveys through 
the pillar to evaluate effectiveness of 
preconditioning. 

6. Fill the haulage drift with ce- 
mented sandfill. An 8:1 sand-to-cement 
dry weight ratio was chosen, based on ex- 
perience in other mining districts. 

7. Drive three-cap raises (manway and 
timber slide, plus two joker chutes) 
through to the haulage drift. Raise 
through cemented fill to provide second- 
ary access to the stopes. Provide air- 
tight raise covers to avoid upsetting the 
ventilation flow. 

8. Leave the present overhand mining 
floor open for access way between stopes. 

9. Begin underhand cut-and-fill min- 
ing by breasting beneath the haulage- 
level cemented fill. Post or timber be- 
neath the haulage floor mat if necessary, 
making the stope as narrow as possible. 
Leave the raises open for secondary es- 
cape as mining progresses downward. 

10. Prepare the floor of each cut with 
caps on 6-ft centers, lagging, wire mesh, 
and Fabrene. Pour cemented fill with 8:1 

^Reference to specific products does 
not imply endorsement by the Bureau of 
Mines. 



53 



Mined and filled 
2 cap raise 2 cap raise 



3 cap raise 



Level 



Sill pillar 



I 




/ 



I Ifcut 11 II II 



nn 



IZL 



3ui 



&^CutJ4_Lr 



T 



-I II 



TTTT 



I 



Undercut and fill stopes 



Mined and filled 



TTTT 



^JaV 



:tft 



Level 



Longitudinal section or undercut and fill stopes 

FIGURE 2. " General layout of pillar recovery plan. 



m 



sand-to-cement ratio (by weight) to a 
depth of 3 to 5 ft, followed by uncement- 
ed tailings for the remainder of each 10- 
ft cut. 



period prior to blasting the destress 
round. A second group of instruments was 
installed during preparation of the de- 
stress round. 



11. Instrument each cut with closure 
extensometers and fill pressure cells to 
monitor load-displacement behavior of the 
fill. 

12. Take the last cut by end slicing. 

Preparations for the underhand cut- 
and-fill experiment began in November 
1980 based on the mining plan. 

INSTRUMENTATION 

The instrumentation program was aimed 
at quantifying the stress and displace- 
ment behavior of the wall rock, ore body, 
and fill before and after the destress 
blasting as well as during subsequent 
mining. 

Initial closure and stress change in- 
struments were installed on the upper 
level during cleanup and repair of the 
September 1980 rock burst damage. These 
were to supply baseline data during 
leveling of the stopes and during the 



Instrumentation performance was mixed. 
The most consistent and useful data were 
obtained from the closure points and 
stope closure extensometers. This clo- 
sure data nicely illustrated the effects 
of preconditioning and pillar behavior 
before and during mining. The stress 
meters provided qualitative data on 
stress conditions in the walls surround- 
ing the test stopes, which could be of 
possible use in a modeling study. How- 
ever, there were too few gauges for a 
quantitative evaluation of the stress 
conditions in the pillar. The soil pres- 
sure cells placed in the fill proved 
disappointing. The gauges are thought 
to have suffered from electronic, envi- 
ronmental, and, possibly, instrument- 
construction problems. Little quantita- 
tive data were produced by the pressure 
cells prior to failure. One point is 
evident from this test: Hand-measurable 
instruments, such as the tape extensom- 
eter, provide the most reliable data in a 
harsh mining environment. 



54 



SITE PREPARATIONS 

Because the mine had no facilities for 
adding cement to the sandfill (mill tail- 
ings), a small cementing system was de- 
signed and constructed on-site. The 
mine's existing sandfill system consisted 
of two major parts (fig. 3): the surface 
sand-pumping facilities located within 
the mill and underground sand storage and 
distribution system. At the mill, a du- 
plex piston pump was used to pump the 
sandfill slurry (40 to 50 wt pet solids) 
through 11,000 ft of 3-in line to the 
2000 level of the mine. At the 2000 
level, the slurry was discharged to a 
cyclone where the underflow was approx- 
imately 53 to 60 pet pulp density and 
contained a high percentage of fines. 
The high percentage of fines bears di- 
rectly on the ability to achieve a good 
sand-cement set. 

After passing through the cyclone, the 
underflow was dumped to a storage tank 
(the storage tank was blasted out of 
country rock and lined with shotcrete) 
and air-agitated. The sandfill was 
gravity-drained from the tank and dis- 
tributed to the required stope through 
3-in black pipe with victualic-style 
couplings. 



Ill Teea 



Gankier-Denver 
duplex piston 
point 



11,000 ft 

3-ln sch 40 pipe 

at 460 pal 



To stopes above 
2000 level 



Gardner-Denver 
-^ duplex piston pump 
at 500 psi 

Overflow — 1 — , 
to-*— I 
tailings I I 

^T^ Cyclone underflow 
1 at 55-60 pet 



Sand storage 



Cement line 
from surface 
plant 



To stopes 
at 90 tons/h 
solids 



2000 level 

FIGURE 3, - Schematic of existing sandfill system. 



A study was conducted to determine the 
most cost-effective method of introducing 
cement into the sandfill for distribution 
to the stopes below the 2000 level. Be- 
cause of operational problems (tramming 
and inability to mix cement in the under- 
ground storage tank) , as well as the cost 
of handling and preparing large tonnages 
of cement underground, the decision was 
made to place the cement slurrying and 
pumping facilities outside the mine. 

The cementing system can be divided in- 
to two component parts: cement handling 
and cement pumping. The components are 
diagramed in figure 4. 

The cement-handling system is that por- 
tion of the cementing system that stores 
bulk cement, delivers it at a desired 
rate, and slurries it to the desired bulk 
density. The functions of the handling 
system are — 

1. Bulk storage of up to 60 dry tons 
of cement . 

2. Delivery of from to 5.5 cfm bulk 
cement for slurrying, with adjustment for 
any range in between. 

3. Capacity for slurrying up to 3,000 
gal of a 50-pct-solids cement slurry. 



Variable speed 
rotary valve feeder 




3,500-gBl 
slurry tank 

Butterfly valve 

Water flush , , ^ 

Accumulator 
Strainer box 



Suction r 



-Dump to tank 



Plug valves 



stabilizer' 
Ash 2-in 
charging pump 

Emergency dump-^ 
to tailings 




2-in mi?\ 
11.000 ft O 



Wilson-Snyder 
43-85T triplex 
plunger pump 



to 2000 

level sandllne 



FIGURE 4, - Schematic of cement-handling system. 



55 



The pumping system referred to here in- 
cludes all pumps, valves, pipelines, and 
accessory equipment for delivering the 
cement slurry to the 2000 level. The 
governing functions of the system are — 

1. A system capable of pumping at a 
rate of 30 to 60 gpm at pressures to 600 
psi during normal operation. 

2. In the event of a need to stop 
pumping into the sandline for short peri- 
ods of time (for example, sandline breaks 
in shaft), the pump should be capable of 
displacing less than 5 gpm of slurry. 

3. The underground line must be capa- 
ble of being flushed both from outside 
with fresh water, or from inside the mine 
in the event of pump failure. 

The design of the cementing system was 
completed in late July of 1980, and or- 
ders were initiated for components in 
August. The initial testing of the sys- 
tem was completed in January 1981. After 
extensive use of the system, the normal 
operating conditions for a slurry pulp 
density of 50 pet were a flow rate of 50 
gpm at a discharge pressure of 450 psi. 

During construction of the cementing 
system, preparation of the floormat in 
the sill drift was completed. When this 
drift was originally driven, 12- by 12-in 
stringers were placed at each rib-floor 
intersection and run parallel to the 
drift axis for its entire length. The 
posts for the drift caps were footed on 
these stringers. The floormat was con- 
structed by placing lagging across the 
stringers and covering the lagging with a 
single layer of 4- by 4-in wire mesh 
(fig. 5) followed by a single layer of 
Fabrene. As no rail track existed, all 
timber was hand-trammed into the drift. 
The rock burst had reduced the drift sec- 
tion to less than 6 ft in height and 
width in certain areas and eliminated 
air and water services. Slabs pulled 
loose from the walls were broken and re- 
moved by hand. Consequently, repairs and 
mat preparation required approximately 
1 month. 




4- by 4-ln No. 8 or doublg 
layer of 6- by 6-ln No. 10 wire 



10- or 12-in lagging 
-12- by 12-ln stringer 



FIGURE 5. - Floormat preparation in over- 
lying drift. 

Shortly after completion of the ce- 
menting system, the drift was filled 
in a series of five individual pours, 
each 100 to 200 ft in length. Owing to 
poor availability of sandfill, this pro- 
cess required approximately 1 month for 
completion. 

By end of January 1981, stope prepara- 
tions were completed. The three stopes 
had been brought to the same elevation 
and were sandfilled (with the exception 
of each raise area) to within 4 ft of the 
back. It was decided to leave this 4-ft 
access way between stopes to provide ease 
of movement during mining and simplify 
preconditioning of the pillar. Each 
raise area was timbered in preparation 
for raise driving from the stope below. 

The original preconditioning design 
called for a series of 35-ft-long, 2- 
1/4-in-diam vertical blastholes drilled 
upward on 6-ft centers the entire length 
of the stope block. Based upon the un- 
derground crew's recommendation, 10-ft 
spacing was adopted, and the holes were 
drilled with jacklegs and stopers using 
1-in rope-thread steel with 2-in and 2- 
1/4-in cross bits. Two to four drillers, 
working on a single-shift basis, required 
less than 2 weeks to drill the 29 holes. 
No holes were drilled at each future 
raise location for fear that blasting 
would create a slabby back. 

Prior to shooting the destress round, a 
seismic velocity survey was performed. 



56 



The data showed P-wave velocities averag- 
ing about 14,000 ft/s. Higher veloci- 
ties, indicating either more competent 
rock, stress concentrations, or both, 
were recorded at the ends of the stope. 
The waves passed through the west area at 
significantly lower velocities. This 
agrees well with the highly broken nature 
of the ore body observed in the stope in 
the vicinity of the September 1980 rock 
burst. In general, the survey velocities 
indicated that both ends of the sill pil- 
lar (not affected by the prior burst) had 
the highest probability of future burst- 
ing. The destress round was loaded and 
shot on March 3, 1981. 

Following preconditioning, the mining 
crews were returned to the three stopes 
and began raising through the sill pil- 
lar. The mining plan called for the 
driving of three-cap raises consisting of 
manway, timber slide, and two joker 
chutes above each previous raise. The 
underground crew, based upon past experi- 
ence, felt that driving three-cap raises 
might open too much ground, increase the 
risk of rock bursting, and create a slab- 
by back. A compromise was reached to 
drive two-cap raises above the two west 
stopes and a three-cap raise in the east 
stope, where conditions were generally 
better. The west stopes would be ser- 
viced from above, and the east stope from 
below. The geometry of the raises and 
the stopes and the preconditioning drill 
pattern are shown in figure 6. 



LavBl 



; I Destress holes jy' Plller jf 



Ik^iilllllllllllll! Ml> 3s>. 



FIGURE 6. • Drilling pattern for pillar preconditioning. 



The driving of the raises proved to be 
fairly difficult in B and C stopes, but 
was accomplished with little difficulty 
in A. The rock burst in the B-raise area 
had left a badly broken ore body and a 
slabby back. C-raise was located in 
fairly competent ground, but an inexperi- 
enced crew overloaded the raise rounds, 
shattering the back and creating a prob- 
lem. Raising required approximately 1 
month in A stope and 1-1/2 months in B 
and C. 

For procedure evaluation, it is noted 
that the miners had some work reserva- 
tions, and absenteeism was a problem. 
Also, the mining method conventionally 
employed at the mine and the particular 
site geometry necessitated some of the 
above-mentioned special preparations. 
The haulage drift above the sill pillar 
had been driven in the vein and, there- 
fore, had to be filled before underhand 
stoping could begin. In addition, raises 
had to be driven through the sill pillar 
because the mine used blind stoping. 
This would not be required where raises 
are developed from level to level before 
stoping commences. 

MINING 

Stoping beneath the sandfill of the 
sill drift began in earnest in May 1981. 
Drifting was accomplished by breasting 
from the raise in each direction using a 
modified V-cut (fig. 7). During the ini- 
tial cut, timber sets were stood beneath 
the floormat because of slimes, old 
track, wood, and loose rock beneath the 
mat. Raise timber required excessive re- 
pairs because of high closure rates. 
This proved to be a continuous problem 
throughout the project and resulted in 
much lost time. Mining the first cut re- 
quired approximately 1-1/2 months in A 
and 2 months in B and C stopes. During 
mining of the first cut, a moderate rock 
burst occurred at some distance out in 
the wall below A and B stopes. Only 
minor damage was seen in the stopes, pri- 
marily loose slabs shaken down in the 
access way at the bottom of the pillar. 



57 



o 

o 
o 

c 
o 

3 

o- 
o 
(0 




-10' 



'I , , ' , t_ 

CROSS SECTION 




LONGITUDINAL SECTION 

FIGURE 7. - Modified V-cut. 



The stopes were prepared for sandfill- 
ing of the first cut. A nominal 1-ft 
layer of broken muck was leveled on the 
floor as a blast cushion, followed by 
heading in of caps on 6-ft centers on top 
of the muck floor. Two rows of lagging 
were nailed down from cap to cap, fol- 
lowed by a layer of 4- by 4-in No. 8 wire 
mesh and Fabrene cloth. The floor caps 
were cabled to the roof caps to prevent 
their slippage and to eliminate the need 
for posting on the cut below. A conven- 
tional sand wall, seen in figure 8, was 
constructed. An initial 8:1 sand-to- 
cement mix was poured to a thickness of 
3 to 4 ft, followed by uncemented sand- 
fill in the upper portion of the cut. 
The high fines content and low pulp den- 
sity of the fill, as well as poor drain- 
age qualities, resulted in a large loss 
of cement down the chutes to the level 



below. Most of the cement and fines re- 
maining in the stope were concentrated 
at the front of the stope near the sand 
wall, leaving an inconsistently cemented 
sandfill over the length of the stope. 

Mining in stope A progressed well dur- 
ing the second cut despite a poorly ce- 
mented sandfill. As shown in figure 9, 
there was no need for additional timber 
in this stope. 

Mining in B and C stopes was combined 
on a double-shift basis, with all broken 
ore slushed to raise B. Inexperienced 
mining crews occasionally blasted down 
some of the overlying caps and fill. 
Raise closure continued at a nearly con- 
stant rate of 1.5 in per month. Raise A 
experienced nearly 18 in of total closure 
during 12 months of mining. The stope 
downtime for replacement of broken raise 
caps was significant, requiring nearly 3 
weeks between cuts. The second cut re- 
quired approximately 1 month to complete 
in stope A and nearly 3 months in B. 

At this time, a complete report of the 
mining and the problem with the cemented 
sandfill was made to the mine management. 
It was agreed that the sandfill crew 
would make a concerted attempt to in- 
crease the pulp density of the fill and 
eliminate the high slimes content. 

Preparation of the floormat was the 
same as in the previous cut, with two ex- 
ceptions: an 18-inch layer of broken 
muck was left on the stope floor as a 
blast cushion, and the caps were tied 
with wire rope to split-set rock bolts 
placed in the wall, as shown in fig- 
ure 10, rather than to the overlying 
caps. Upon filling, however, problems 
with low pulp density and high slimes 
content were again encountered, and poor 
sandwall and stope drainage technique 
caused a high cement loss down the 
chutes. 

The mining of the third cut in stope A 
went quite well. Experience, gained in 
the previous cuts, eliminated basic oper- 
ational problems. Once mining had pro- 
gressed two rounds on either side from 



58 




FIGURE 8. - Sand wall construction. 



the raise, it was possible to cycle each 
stope face almost nightly. Tying the 
caps to split-set bolts with wire rope 
eliminated the need for additional tim- 
ber. The marginal cemented fill in this 
cut presented no significant problems 
due to the narrow (6 to 7 ft) width of 
the stope. Mining of this third cut 
required approximately 1 month, includ- 
ing 1 week of raise repair prior to min- 
ing. Productivity and cost were greatly 
improved. 

Closure of the stope continued at ap- 
proximately the same rate as before. 
Low mlcroseismic activity was likely re- 
lated to frictional sliding along frac- 
tures in the crushing pillar. The data 
indicated the pillar was failing in a 



nonviolent manner with 
of blasting. 



little likelihood 



During this time, the second cut in 
stope B was completed and prepared for 
fill. During filling, greater attention 
was directed to proper drainage, result- 
ing in less spillage of cement to the 
level below. This produced the best fill 
to date, as was observed when mining be- 
gan in the next cut ,(fig. 11). Though 
still not of the quality desired, the 
fill was strong enough to support itself. 
Both stopes were experiencing a continual 
problem with much removal. Prior cement 
spills , which had reached the pockets on 
the lower level, cemented the broken ore. 
The result was inadequate storage capa- 
city and muckbound stopes. 



59 




\ ■■ V • 



V 1 '. 



FIGURE 9. - Completion of mining in second cut. 



60 





FIGURE 10. - Stope prepared for filling. 



Preparation of the third-floor mat in 
stope A required approximately 1 week. 
The mining crew handled the sandfilling 
duties themselves. They made an initial 
35-min pour, then held the water behind 
the sand wall, allowing the cement to 
settle out. The water was then decanted 
off, and the remainder of the stope was 
filled. A fairly well-cemented fill was 
obtained. By the end of December 1981 
mining was completed on the third cut of 
stope B without major mining problems. 
The hung ore pocket continued to cause a 
muck removal problem. 

The fourth and final cut, which was 14 
ft in height, was mined by end slicing, 



but with conventional br easting-down 
rounds rather than the horizontal V-cut 
used thus far. The mining in stope A 
progressed rapidly and was completed by 
mid-March. This crew had become quite 
proficient in underhand mining and had 
few problems over the last two cuts. 

In this final cut, a decrease in clo- 
sure rate was reflected by fill pressure 
of approximately 100 psi (before erratic 
readings developed) . Slow progress was 
made in stope B because of many boulders, 
which restricted the access way at the 
bottom of the pillar. Poor ore prices 
caused layoffs at the mine, and the 
crews were changed in B, further slowing 



61 



^^^ ' 



m;r 



FIGURE 11. . Cement fill above third cut. 



progress. A general mine closure even- 
tually halted mining at the end of May 
1982, with an approximate 50-ft length 
of pillar remaining. Mining of the pil- 
lar following initial raising had re- 
quired 12 months in stope A and some 
14 months in B. 

COST COMPARISON 

For cost analysis during the demonstra- 
tion phase, the actual production costs 
are given in table 1. Data comparison is 
made to mining of a 35-ft-thick sill 



pillar in a nearby area of the mine by 
the overhand method. 

It is estimated that underhand cut- 
and-fill costs would be $48.81 per ton If 
the method was adopted as a standard pro- 
cedure for mining sill pillars. This 
estimate argues that experienced crews 
in well-prepared stopes would reduce 
costs in the areas where unscheduled 
maintenance proved costly in terms of 
lost production, as well as extra labor 
and materials. 



62 



TABLE 1. - Comparison of sill pillar recovery costs, per ton 



Item 



Labor 

Explosives 

Rock bolts , 

Sandf ill 

Drill repair and drills , 

Bits and rods 

Timber 

General-conventional-miscellaneous, 

under hand-Fabrene , cement 

Total direct costs per ton , 

Total indirect costs per ton , 

Total costs (excluding milling and 
general and administrative) 

Productivity. .. .tons per worker shift., 



Overhand 

cut and 

fill 



$19.94 
1.54 
1.81 
1.25 
1.01 
.77 
1.97 

1.99 



30.28 
27.08 



57.36 



12.04 



Underhand 

cut and 

fill 



$22.56 

1.15 

.25 

1.42 

.95 

.91 

2.09 

6.68 



36.01 
27.41 



63.42 



7.96 



CONCLUSIONS 



Little difficulty was experienced in 
applying underhand cut-and-fill mining 
procedures to the Coeur d'Alene mines. 
After initial problems were resolved, 
rounds were cycled daily in the later 
cuts. Two noteworthy problems were en- 
countered, however. Excessive closure in 
the raises resulted in a great deal of 
raise repair. Carefully designed and 
constructed raises would save considera- 
ble time and expense in light of the 
amount of closure that can be expected in 
a preconditioned pillar. The second 
problem was getting a well-consolidated 
cemented fill. Particle size of the mill 
tailings was very fine and detrimental to 
setting of the sand-cement mix. Also, 
the weight percent solids of the mix was 
too low. Excess water prevented cement 
from adhering to the sand particles, al- 
lowing it to be carried off with the 
slimes. Control of the pulp density is 
critical and should be kept at 65 pet 
minimum. 

Regardless of operational problems, the 
success of the project must be measured 
in terms of rock burst reduction. The 
only rock burst occurred some 100 ft out 
in the wall below stopes A and B (during 
mining of the first cut), causing minimal 
damage in the undercut stopes. Five caps 



in stope B were broken, but it was not 
apparent whether this was caused by the 
burst or by the 4 in of closure that re- 
sulted from mining. A minor amount of 
rock slid off the walls. Only 0.3 In of 
closure was due to the burst. In the old 
stopes below, considerable rock was 
shaken from the back. There was no 
seismic buildup prior to the burst, and 
the popping and cracking accompanying 
mining was of a destressing nature, too 
small to register on the mine micro- 
seismic monitoring system. 

The last three cuts were mined without 
the occurrence of bumps or rock bursts. 
The minor popping and cracking that 
accompanied mining of the first three 
cuts disappeared by the fourth (final) 
cut as the stopes continued to squeeze 
and further destress. The high closure 
accompanying the mining of all cuts , 2 
to 4 in per cut, verified the effective- 
ness of destress blasting in softening 
the pillar. 

Underhand mining beneath a cemented 
backfill appears to offer greatly im- 
proved ground control during sill pillar 
mining. Provided the cemented fill is of 
good quality, there is no potential roof 
fall problem. 



63 



A comparison shows the cost of under- 
hand cut and fill was 11 pet higher and 
productivity was 34 pet lower than that 
associated with mining a similar pillar 
using conventional overhand cut and fill. 
This is largely due to the mine's inex- 
perience with the mining system (includ- 
ing preparation) and cemented backfill. 
If the costs of the underhand cut and 
fill are adjusted to reflect an accepta- 
ble productivity level, the underhand 



cut-and-fill mining will reduce sill pil- 
lar mining costs by 15 pet. 

The conclusion from this demonstration 
is that the combination of ore precondi- 
tioning and underhand mining resulted in 
greatly improved rock burst and ground 
control, allowing safe and efficient min- 
ing of a potentially hazardous, burst- 
prone pillar. 



64 



HAZARD DETECTION 

ACOUSTIC CROSS-BOREHOLE SYSTEM FOR HAZARD DETECTION 

By Karen S. Radcliffe^ and Richard E. Thill2 



ABSTRACT 



A high-frequency (20 kHz) acoustic 
cross-borehole system has been devised 
and field tested to remotely investigate 
the structural conditions of a rock mass 
in advance of mining. Elastic properties 
of the rock can be determined by monitor- 
ing acoustic waves generated between 
boreholes, and the structural integrity 
of the rock mass can be interpreted by 



evaluation of the acoustic signal charac- 
teristics. Identification of hazardous 
ground conditions prior to mining can 
help reduce, and ultimately prevent, in- 
juries and fatalities to miners as well 
as interruptions in the mining operation 
due to encounters with unstable ground 
and geologic hazards. 



INTRODUCTION 



Geologic anomalies can have serious 
effects on the safety of miners and on 
production when encountered unexpectedly 
during a mining operation. Unidentified 
fracture zones, voids (such as abandoned 
mine workings or solution cavities) , 
lithologic f acies changes , and inclu- 
sions, both within and surrounding the 
ore body, create potentially hazardous 
conditions for the miner. Fractures, 
joints, faults, and bedding planes are 
also critical in determining environmen- 
tal effects from mining, controlling 
ground movements, and supporting excava- 
tions and mine structures. 

These structural features create zones 
of increased permeability, often con- 
taining large quantities of ground water 
that can quickly inundate the mine when 
encountered during mining. These same 
structural features can also produce 
weakened and therefore unstable roof 
conditions, resulting in roof falls or 
collapse at the working face. Also haz- 
ardous in coal mining is the methane 
that may be contained in fractures or 
joints. Abandoned and unmapped oil and 
gas wells, and especially abandoned mine 
workings, create vulnerable conditions 



Geologist. 



^Supervisory geophysicist. 
Twin Cities Research Center, Bureau of 
Mines, Minneapolis, MN. 



for inundation and mine 
cially in coal mines. 



flooding, espe- 



In addition to the structural weakness 
inherent in a rock mass from natural dis- 
continuities and geologic features, dam- 
age can also be introduced into the 
mine's supporting structures from mine 
excavation and blasting operations. Even 
controlled blasting can result in over- 
break into mine structures or weaken al- 
ready unstable areas. 

To help reduce, and ultimately prevent, 
injuries and fatalities to miners and 
interruptions in the mining operation 
from unstable ground and other geologic 
hazards, it is necessary to remotely in- 
vestigate the structural conditions in 
advance of mining. This can be accom- 
plished on a large scale during mine ex- 
ploration and development, or on a small 
scale as mining progresses by probing the 
rock mass structure ahead of the working 
face. It is necessary to characterize 
the nature and structural integrity of 
the rock mass surrounding the material to 
be mined as well as the integrity of the 
ore body itself. By characterizing these 
materials and identifying any hazardous 
anomalies prior to rock excavation, ap- 
propriate safeguards can be taken to re- 
duce or prevent the hazards posed by un- 
stable conditions. 



65 



Various techniques exist for assessing 
the condition of a rock mass , ranging 
from mechanical property tests on small- 
scale samples in the laboratory to large- 
scale testing under field conditions. 
Although laboratory testing provides val- 
uable information concerning the intact 
elements of the rock mass, the effects of 
large-scale discontinuities in that rock 
mass are best determined in situ. Me- 
chanical property data that are truly in- 
dicative of the behavior of the rock mass 
under in situ conditions of stress, mois- 
ture, and other environmental factors are 
also best determined in the field. In 
situ techniques also provide the only op- 
portunity to identify and specifically 
locate the presence of geologic anomalies 
and hazardous conditions. 

Depending upon the type and size of 
feature to be located, several in situ 
geophysical methods for evaluating the 
structure of a rock mass are available. 
Nonseismic geophysical surveying includes 
gravimetric, magnetometric and geoelec- 
tric, thermal, radioactive, and geochemi- 
cal methods (J_).2 Application of these 
techniques, as well as seismic methods, 
is based on the presence of a measurable 
difference in a physical property within 
the earth materials under evaluation, 
such as acoustic velocity, density, mag- 
netic susceptibility, or resistivity. 
Use of nonseismic geophysics in detecting 
mining hazards is limited by the range 
over which the methods can be applied and 
by the degree of variation in physical 
properties required within the structures 
of interest. Interpretation of field 
data is also limited because of its de- 
pendence upon the model chosen in the 
investigation. 

ACOUSTIC CROSS 

The acoustic cross-borehole system op- 
erates at a frequency of 20 kHz. It cou- 
ples under pneumatic or hydraulic pres- 
sure to the borehole wall, and can there- 
fore function in water-saturated or dry 
holes. Acoustic measurements can be made 
in vertical holes from the surface, or in 



■^Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this paper. 



Seismic techniques have been used ex- 
tensively by the petroleum industry for 
locating geologic structures favorable 
for economic accumulations of oil and 
gas. Surface seismic exploration methods 
have also been used for determining sub- 
surface geology and mapping. Of all the 
physical methods used in geological ex- 
ploration, the seismic methods are con- 
sidered to be the most direct, and when 
applicable, give the least ambiguous 
results. Most of the techniques used are 
surface surveys that operate over a large 
horizontal distance. The vertical depth 
to which seismic surveys are effective 
depends upon the stratigraphy and struc- 
ture as well as the strength and fre- 
quency of the seismic signal utilized in 
the survey. The applicability of seismic 
methods is based on the relationship 
between the acoustic properties of rocks 
and their physical properties , mineral 
composition, and structural integrity 
(2). 

To overcome the deficiencies of many 
surface geophysical methods, an acoustic 
cross-borehole device was developed and 
field tested by the Bureau of Mines (3^) . 
The apparatus is designed to operate at 
high frequency (20 kHz) over moderate 
distances between boreholes. These dis- 
tances range from a few meters to tens of 
meters, depending upon the acoustic 
transmission characteristics of the rock. 
Applications include evaluating the elas- 
tic properties and integrity of under- 
ground structures, monitoring stress 
changes in and around underground work- 
ings, evaluating fragmentation and over- 
break from blasting, and detecting dis- 
continuities in advance of mining. 

-BOREHOLE SYSTEM 



horizontal or inclined holes in the roof, 
rib, or floor of an underground mine. 

SYSTEM COMPONENTS 

The basic acoustic cross-borehole sys- 
tem consists of a high-frequency pulse 
generator, transmitting and receiving 
borehole transducers, and monitoring and 

timing electronics (fig. 1). Auxililary 
components include a downhole amplifier 
in the receiver transducer to improve the 



66 




FIGURE 1. - Basic components of acoustic cross-borehole system: borehole transducers, pulse gen- 
erator, and signal monitoring and timing electronics. 



received acoustic signal, and high- 
voltage coaxial cable for driving the 
source. 

Pulse Generator 

A 5-kV electronic pulse generator 
drives a piezoelectric source in the 
downhole transmitter. A fast-rising out- 
put trigger pulse is provided for syn- 
chronizing the timing circuitry. The 
unit is operable from a conventional 60- 
Hz, 120-V line source. 

Borehole Transducers 

The transmitter and receiver transduc- 
ers are pressure-sensitive hydrophones. 
Uniform lateral radiation and reception 
of acoustic energy is achieved through 
use of cylindrical lead-zirconate piezo- 
electric transducers operating in the ra- 
dial expansion mode. A cross-sectional 
view of the transducer assembly is shown 
in figure 2. An internal inflatable 
bladder causes displacement of fluid in 
the transducer cavity and outward expan- 
sion of the outer neoprene boot to cou- 
ple to the borehole wall. The transducer 



elements are moderately damped, giving 
high sensitivity but producing some re- 
verberation ringing. The reverberation 
is not detrimental to picking the onset 
of the first arrival normally associated 
with the compressional wave (P-wave) , but 
it can mask or interfere with subsequent 
wave arrivals. The emitter and receiver 
transducers are identical in construction 
and are 6.67 cm diam by 20 cm long. The 
central operating frequency of the emit- 
ter is 20 kHz. Frequency response of 
the receiver is essentially flat between 
19 and 31 kHz. Operating voltage for 
the emitter is limited to 2 kV to pre- 
vent deterioration of the piezoelectric 
elements. 

Downhole Amplifier 

A solid-state amplifier was designed to 
fit into a small (7-cm-diam) compartment 
located near the receiver transducer. 
The amplifier provides substantial sig- 
nal gain and drives a long length of 
cable (up to 25 m) with minimum signal 
loss or distortion. Frequency response 
of the amplifier is flat over the range 
5 to 50 kHz. 



67 



Pressurizin 
port 

Inflatable 
bladder 

Stainless 

steel 

housing 

Couplant 



Neoprene — ^ 

expandable 

boot 

Electrical 
insulator 




Coaxial 
connector 

Air or fluid 

pressurizing 

media 



Electrode 
feedthrough 



Cylindrical 
piezoelectric 
transducer 
element 



Transducer 
backing 

Boot clamp 



FIGURE 2. - Cross-sectional diagram of bore- 
hole transducer assembly. 

Monitoring and Timing Circuitry 

Conventional oscilloscope waveform dis- 
play and timing circuitry are used to 
monitor the received signal and to obtain 
elastic wave transmit times. Acoustic 
signals are captured and digitized on a 
floppy disk to permit detailed waveform 
analysis. 

PRINCIPLES OF OPERATION 

The principle of the acoustic cross- 
borehole system is to propagate a seismic 
wave from an emitter in one borehole to a 
receiver in a second borehole (fig. 3). 
By measuring the wave's traveltime and 
the distance between boreholes, the 
acoustic velocity of the seismic wave 
through the rock mass can be determined. 
By moving the emitter and receiver along 
the lengths of their boreholes, profiles 
of wave velocity or amplitude in the 



plane including the boreholes can be 
made. Proper geometric arrangement of 
the boreholes then enables delineation of 
the subsurface structures. Calculation 
of wave velocity assumes a straight line 
of propagation between the emitter and 
receiver transducers, and for short prop- 
agation paths, the direct compressional 
wave is normally the first energy to ar- 
rive at the receiver. At greater dis- 
tances, however, and in the presence of 
higher velocity zones adjacent to the 
path between the emitter and receiver, 
refracted waves may begin to reach the 
receiver before the direct P-wave. Typi- 
cal wave traces are shown in figure 4. 

Deviations from the expected character- 
istics of the propagated wave can be used 
to identify structurally hazardous fea- 
tures in mine strata. Characteristics of 
the seismic wave arriving at the receiver 
(such as the wave velocity, amplitude, 
and frequency content) are related to the 
physical and mechanical properties of the 
geologic media through which they travel. 
In rock masses approximating homogeneous, 
elastic, isotropic media of semi-infinite 
extent, a seismic disturbance will gen- 
erate compressional (P-wave) and shear 
(S-wave) waves that travel with veloci- 
ties related to the density and elastic 
constants of the rock medium according to 
the equations 



and 



"P [p (1 + a)(l-2a)J 



/2 



(1) 
(2) 



where Vp is the compressional wave veloc- 
ity, Vs is the shear wave velocity, p is 
the density, a is Poisson's ratio, and E 
and G are the Young's and shear moduli, 
respectively. Surface waves may also be 
generated at free surfaces of the rock 
mass by the seismic disturbance (_3 ) . 

Unfortunately, rock masses are seldom 
homogeneous, perfectly elastic, or iso- 
tropic, but are more often complex struc- 
tures containing joints, bedding, frac- 
tures, and weathered zones. Discon- 
tinuities such as these, in addition to 



68 



— il^Surface support system 




^^T^^^^^J^^^JT-f- 



\t='y^^^ 



Emitter 




Zone of Discontinuities 



FIGURE 3. - Acoustic system for detecting discontinuities between boreholes. 



cavities and voids within the rock mass, 
all influence the transmission of seismic 
waves through the rock. The presence of 
structural features in a rock mass can be 
identified by monitoring deviations in 
the compressional- and shear-wave veloci- 
ties as well as the associated mechanical 
properties that can be calculated. The 
structural integrity of a rock mass is 
assessed by comparing the wave velocity, 
wave amplitude, other signal characteris- 
tics, and the physical and mechanical 
properties of the rock mass with known 
characteristics of the intact material. 

Reliable detection of geologic hazards 
requires that there be sufficient acous- 
tic contrast between the target hazard 
and the surrounding material so that the 
probing energy will be reflected, re- 
fracted, attenuated, or otherwise identi- 
fiably modified. Also important with 
respect to the frequency of the seismic 
signal generated are the target size and 
shape compared with the wavelength of the 



signal. The ability to accurately inter- 
pret information about geologic condi- 
tions from the detection signals requires 
prior identification of various types of 
geologic anomalies and what their effects 
on propagated seismic signals might be. 

Propagating and detecting seismic waves 
in a discontinuous rock mass requires a 
tradeoff between range and resolution. 
Use of high-frequency seismic energy re- 
sults in better resolution of small geo- 
logic features but limits the penetration 
distance, because of the selective atten- 
uation and absorption of the higher fre- 
quency components of the propagating 
wave. Seismic resolution of geologic 
anomalies that may cause mining hazards 
requires use of a wavelength comparable 
to, and preferably smaller than, the 
dimensions of the anomaly to be observed. 
Table 1 describes typical applications 
of the cross-borehole acoustic technique 
relative to the seismic frequencies 
required. 



69 



157 cm 



319 cm 



473 cm 




603 cm 



Wiiiil 






I 



.5 V 



h 






0.5 ms 

FIGURE 4. - Typical wave traces obtained in situ. 

TABLE 1. - Typical applications of acoustic cross-borehole methods 



Operational frequency 


20 kHz 


2 kHz 


Typical wavelength in rock, m: ' 






Sandstone. ................... 


0.10-0.18 

.16- .28 

.11 

.06- .14 

Fractures, joints. 


1.0-1.8 


Limestone. ................... 


1.6-2.8 


Shale 


1.1 


Coal 


.6-1.4 


Geologic hazards or structural 


Faults , channel 


features detected. ^ 


kettlebottoms, 


sands, voids. 




blast damage. 


abandoned mines , 




small-scale 


large-scale 




geologic dis- 


geologic dis- 




continuities. 


continuities. 



'Reference 2. 

^Rock mass properties (Young's modulus, shear modulus, bulk modulus, 
Poisson's ratio) can also be determined at each frequency for various 
rock types. 



70 



PROOF TESTING AND LABORATORY CALIBRATION 



The performance capability of the 
acoustic cross-borehole apparatus was 
first tested in a block of granite in the 
laboratory. Tests were conducted to de- 
termine waveform characteristics, P-wave 
velocity, the radiation pattern from the 
source, the traveltime correction factor 
for electronic delays and pulse buildup 
in the transducer , and the optimum cou- 
pling pressure. 

Testing determined that a minimum pres- 
sure of 10 psi was required to couple the 
transducer to the borehole wall and that 
amplitude of the first arrival gradually 
increased until 50 psi, after which there 
was little change. All subsequent tests 
were therefore conducted at coupling 
pressures between 50 and 70 psi. A near- 
ly uniform radiation pattern was observed 
by monitoring the output from an accel- 
erometer as the receiver was rotated 
around the source. Small irregularities 
were attributed to the coupling and sur- 
face conditions. Wave transmission tests 
produced excellent signal reception with 
well-developed first arrivals. Rise time 
of the first arrival was about 5 us, 
indicating that high frequencies were 



retained over short transmission dis- 
tances in the granite. This permitted 
wave velocity to be determined within 1 
pet precision (_3) . 

Mechanical properties of the block were 
established in advance by conventional 
ultrasonic and bar resonance tests (4_) on 
core extracted from boreholes in the 
granite block. These properties provided 
a basis for comparison of the laboratory 
data with in situ measurements. 

Laboratory testing was also performed 
in a concrete slab to determine the ef- 
fects of propagation distance on the am- 
plitude of the first wave arrival. The 
wave traces verified that the first arri- 
val diminishes rapidly with increasing 
propagation distance. As demonstrated 
in the granite block, the high frequen- 
cies were filtered through the concrete. 
Although the time difference between 
first and second arrivals increases with 
propagation distance, the change is not 
large and again introduces an error of 
less than 1 pet in determining transit 
time at these distances. 



FIELD VERIFICATION 



EMERALD ISLE MINESITE 

The first field tests of the high- 
frequency cross-borehole system were con- 
ducted at the Emerald Isle open pit mine 
near Kingman, AZ, where the Bureau of 
Mines was experimenting with blasting to 
fracture the rock mass for in situ leach- 
ing {5). Acoustic velocities of a cop- 
per conglomerate were determined in deep 
(76 to 85 m) , vertical, water-saturated 
boreholes following blasting. The tests 
were designed to determine whether the 
cross-borehole system could be used to 
detect increased fracturing following 
blasting. Results showed a decrease in 



acoustic velocity with depth, correspond- 
ing to a deterioration of the rock due to 
blasting, and delineation of a weathered 
zone at the granite-gneiss interface. 
Wave velocities calculated from field 
data are shown in relation to velocities 
determined from laboratory tests on pre- 
shot core recovered from the boreholes 
(fig. 5). 

SENECA MINESITE 

Performance tests of the cross-borehole 
system were also conducted in inclined, 
water-filled boreholes underground at 
the Seneca Mine near Mohawk, MI. In 



71 



175 



205 



t235 



265- 



295 



KEY 
^ Cross-borehole velocity 

determinations 
— Preshot P-wave velocity 

Overburden 
contact I 



\ Gila conglom- 
erate 



A Twoaiiioiou &UIIO 



RubbI 



^^ 



Weatheredzone 

Granite gneiss 
contract 



O A §1 

MEAN VELOCITY, km/s 



8 



60 



70 E 

I 
I- 

Q. 
LU 

Q 
80 



90 



FIGURE 5. - Comparison of laboratory and field 
acoustic velocities at the Emerald Isle minesite. 



Studies similar to those at the Emerald 
Isle site, the Bureau of Mines was con- 
ducting research to develop methods of 
confined blasting to create fracture 
permeability for in situ leaching of na- 
tive copper ores (6^). Acoustic veloci- 
ties were determined before and after 
blasting at a separation distance of 5m 
between boreholes. The velocity deter- 
minations indicated little, if any, ef- 
fect from blasting in the zone of the 
explosives column. Results from permea- 
bility and Rock. Quality Designation (RQD) 
measurements conducted at the same time 
also indicated that the rock was not 
highly fractured by the blast and that 
fractures which were present were tightly 
closed by overburden pressure O) . The 
excellent quality of the signals received 
indicated that considerably larger dis- 
tances (>15 m) could have been traversed 
in the amygdaloidal basalt host rock. 



LOGAN WASH MINESITE 

The most recent series of field tests 
was conducted in cooperation with 
Occidental Oil Shale, Inc., at the Logan 
Wash experimental mine near DeBeque, CO. 
Acoustic velocities were determined under 
preshot and postshot conditions as in 
situ oil shale retorts were rubblized us- 
ing explosives. The field tests were 
designed to evaluate the cross-borehole 
system's capability to detect blast dam- 
age to the pillars separating the re- 
torts. One set of measurements was taken 
in vertical holes, and two other series 
were made at inclinations of 30° and 45° 
in water-saturated boreholes. 

Slight decreases in acoustic velocities 
were observed following retort rubbliza- 
tion, suggesting increased fracturing in 
the pillars (fig. 6). The greatest over- 
all decrease in velocity was on the order 
of 4.5 pet in a zone between 12 to 15 m 
from the borehole collar and 14 m in from 
the retort wall. It was expected that 
there would be a minimal change in veloc- 
ity before and after blasting, since the 
measurements were made in the pillar and 
the blasting was designed to contain 
fragmentation within the rubble bed in 
the retort. 

Cores extracted from the pillar be- 
fore and after retort rubblization were 
subjected to ultrasonic velocity deter- 
mination in the laboratory to provide 
another set of data for detecting veloc- 
ity changes. The observed decrease in 
field velocity following blasting cor- 
responds favorably with the laboratory 
acoustic data, as well as with RQD and 
fracture frequency determined from the 
preshot and postshot core. This posi- 
tive correlation suggests that the cross- 
borehole system is a reliable tool for 
detecting small variations in geologic 
structure in situ. 



72 



E 



u 
o 

ui 

> 

lU 

(9 

< 
oc 
IIJ. 



5.50 



5.25 



5.00 



4.75 



4.50 



4.25 



4.00 



3.75 



3.50 




KEY 

Preshot R-8 
Postshot R-8 



10 



15 



20 



25 



45 



50 



55 



60 



65 



30 35 40 

DEPTH, ft 

FIGURE 6. - Depth versus acoustic velocity, preshot versus postshot Retort 8, Logan Wash site. 
DEVELOPMENT OF LOW-FREQUENCY CROSS-BOREHOLE SYSTEM 



70 



A second cross-borehole system is cur- 
rently in development by the Bureau of 
Mines that operates at lower frequency (1 
to 2 kHz) over wider separation distances 
between boreholes. The low-frequency 
seismic signal generated by this system 
will provide complementary hazard detec- 
tion capabilities to the high-frequency 
system by detecting much larger anomalous 
features prior to mining, and at dis- 
tances of up to 150 m. This system is 
especially designed to detect uncharted 
underground voids in coal measure rocks. 
It is unique in that it will operate 
simultaneously in combined modes of re- 
flection and through-transmission to pro- 
vide the best opportunity to detect haz- 
ards in advance of mining (fig. 7). 

The acoustic principles utilized by 
this low-frequency system are the same as 
those governing the operation of the 



high-frequency cross-borehole system cur- 
rently in use. The combined reflection, 
through-transmission capability, however, 
provides for a modified method of struc- 
ture interpretation. In this case, the 
emitter probe located in one borehole 
sends out an acoustic pulse toward the 
receiver probe in a second borehole. A 
void space, or other geologic discon- 
tinuity, will be delineated by the record 
of transmitted and reflected energy in 
the acoustic waveform generated by the 
emitter. A sharp discontinuity between 
boreholes will reflect back nearly all 
energy to the receiver-emitter probe and 
transmit little energy to the distant 
receiver. Distance to the discontinu- 
ous surface can be calculated from the 
traveltime data. Conversely, in a solid, 
homogeneous rock mass, nearly all of the 
energy will be transmitted and little re- 
flected. By profiling along the lengths 



73 



Surface support 
system and vehicle 




o^ 



\x 



.,-m^^-'- -'-' 



<??? 




'^r^^'??^^;?^!^^^::^^^. 



'"mw^m%' 



Transmitter- 
receiver probe 



Acoustic 

source 



Acoustic 
receiver 



Abandoned mine 




FIGURE 7. - Combined reflection and through-transmission cross-borehole system. 



of the boreholes , and by arranging geo- 
metric patterns of boreholes, voids and 
other discontinuities could be delineated 
in size, shape, and location. Comparison 
of the transmitted and reflected energy 



will be useful in interpreting hazards 
ahead of mining. Sophisticated data pro- 
cessing techniques will also permit more 
detailed reduction of velocity and fre- 
quency data. 



CONCLUSIONS 



A high-frequency acoustic cross- 
borehole system has been developed and 
field tested by the Bureau of Mines for 
in situ investigation of rock mass prop- 
erties and detection of geologic hazards 
in advance of mining. The system oper- 
ates at a frequency of 20 kHz between 
boreholes separated by distances from a 
few meters to tens of meters. Changes in 
the characteristics of a seismic wave as 
it propagates through geologic structures 
can be used to detect potential hazards 
in the mining environment. Applications 
of the cross-borehole system include de- 
tection of fracturing and overbreak from 
mine blasting, locating underground voids 
or abandoned mine workings , inferring 
stress distributions in mine structures. 



and detecting geologic hazards in advance 
of mining. 

The second cross-borehole system cur- 
rently under development will propagate 
much lower frequency (1 to 2 kHz) seismic 
waves between boreholes separated by up 
to 150 m. This system will operate in a 
combined reflection, through-transmission 
mode to detect and define the size, 
shape, and location of larger geologic 
hazards in advance of mining. The system 
is specially designed to detect uncharted 
voids in coal measure rocks. Field test- 
ing of the low-frequency system will fol- 
low assembly of the equipment components, 
currently in progress. 



74 



REFERENCES 



1. Dohr, G. Applied Geophysics. Ge- 
ology of Petroleum, v. 1. Halsted Press, 
New York, 1981, 231 pp. 

2. Rzhevsky, V., and G. Novik. The 
Physics of Rocks. MIR Publ. , Moscow, 
1971, 320 pp. 

3. Thill, R. E. Acoustic Cross-Bore- 
hole Apparatus for Determining In Situ 
Elastic Properties and Structural Integ- 
rity of Rock Masses. Paper in Proc. 19th 
Symp. on Rock Mechanics, Stateline, NV, 
May 1-3, 1978. Univ. NV, Reno Press, 
pp. 121-129. 

4. Thill, R. E., and S. S. Peng. Sta- 
tistical Comparison of the Pulse and 



Resonance Methods for Determining Elastic 
Moduli. BuMines RI 7831, 1974, 24 pp. 

5. D'Andrea, D. V., W. C. Larson, 
L. R. Fletcher, P. G. Chamberlain, and 
W. H. Engelmann. In Situ Leaching Re- 
search in a Copper Deposit at the Emer- 
ald Isle Mine. BuMines RI 8236, 1977, 
43 pp. 

6. Chamberlain, P. G. In-Place Leach- 
ing at the Seneca Mine, Mohawk, Michigan. 
Paper in Upper Peninsula AIME Spring 
Technical Meeting Review. Skillings' 
Min. Rev., v. 66, No. 21, May 21, 1977, 
pp. 19-20. 



75 



IN-SEAM GEOPHYSICAL TECHNIQUES FOR COAL MINE HAZARD DETECTION 
By Richard J. Leckenby^ and James J. Snodgrass^ 

ABSTRACT 



The Bureau of Mines has helped to make 
a number of advances in improving and de- 
veloping geophysical methods for detect- 
ing mining hazards from the working face 
in underground coal mines. This paper 
presents the results of some of the Bu- 
reau's recent work in four, in-seam geo- 
physical methods: 

1. A synthetic pulse radar system with 
initial field test results showing its 
potential. 



2. A high-frequency prototype seismic 
system used to ascertain development cri- 
teria for a hand-held scanner. 

3. Controlled-source piezoelectric 
transducers used to generate predominant- 
ly compressional or shear wave energy for 
better control over propagation of guided 
waves . 

4. A dry-hole borehole sonic probe 
adapted to determine compressional and 
shear wave velocities in a coal seam from 
a horizontal borehole. 



INTRODUCTION 



Coal mines are often plagued by adverse 
ground conditions because of geology or 
previous mining or drilling in the area. 
Common geologic problems include faults , 
channel sands, split seams, and clay 
veins , while manmade problems include 
abandoned mines and wells. These prob- 
lems can cause roof falls , blowouts , 
flooding, or inundation. It is important 
to determine the existence and locations 
of these problem areas so that precau- 
tions can be taken to correct or avoid 
potential hazards. In addition, timely 
detection can lead to increased produc- 
tivity and resource recovery. 

One of the most common and costly meth- 
ods of hazard detection is drilling, but 
even extensive drilling programs can miss 

^Physicist. 
^Geophysicist. 
Denver Research Center, Bureau of 
Mines, Denver, CO. 



significant geologic and manmade con- 
ditions. The Bureau has been investi- 
gating, devising, and improving in-seam 
geophysical methods for accurate, relia- 
ble, economic, and practical detection 
and mapping of potential hazards from the 
working face. 

There is no one proven method that can 
detect all hazards in all conditions and 
locations. It is necessary to undertake 
research to devise a variety of methods 
that can be used either singly or in 
combination to provide for the most 
efficient, reliable, and accurate hazard 
detection. This paper considers some of 
the most recent technological advances 
the Bureau has made in in-seam geophysi- 
cal methods. It includes a new ground 
probing radar system called "synthetic 
pulse radar" and seismic methods which 
include developments in high-frequency, 
seismic, guided waves and borehole 
techniques. 



76 



SYNTHETIC PULSE RADAR 



Ground probing radar is a method using 
electromagnetic waves for the detection 
and mapping of anomalous conditions in 
the ground. A number of ground probing 
radar methods are either being used or 
being investigated for use.^ Short pulse 
radar is probably the most commonly used 
method, but electronically it is diffi- 
cult to generate a high-power, high- 
frequency, broadband short pulse that can 
be used to detect in-seam coal mine haz- 
ards up to 200 ft from the working face. 
Synthetic pulse radar is a method that 
overcomes some of the difficulties found 
in short pulse systems. 

Synthetic pulse radar is based on a 
proprietary concept developed by ENSCO, 
Inc. A feasibility study was conducted 
first under a Bureau contract to deter- 
mine the potential of the concept for im- 
proving radar technology.^ Based on that 
first study a cost-sharing contract was 
awarded to XADAR Corp. (a subsidiary of 
ENSCO, Inc.) to design, build, and test a 
prototype system for use in underground 
coal mines. 



short pulse radar, it is useful to review 
some of the basic principles behind 
Fourier analysis. According to Fourier 
theorem, any periodic function (wave) can 
be represented as a sum of a number (pos- 
sibly infinite) of sine and cosine func- 
tions. That is, a periodic wave can be 
given by the equation: 

Y = ao + ai sin tot + a2 sin 2ajt 

+ a3 sin 3tot + . . . 

ai'cos cot + a2'cos 2(jot 

+ a3'cos 3a)t + ... (1) 

This equation is known as a Fourier 
series, which is composed of a constant 
ao , amplitudes a], a2 , a.-^,,,. a^", a2'', 
a3' ... and angular frequencies o), 2a), 
3oj, ... The resultant wave is regarded 
as being built up by a number of waves 
whose wavelength ratios are 1, 1/2, 1/3, 
1/4, ... From the study of sound this 
represents the fundamental and its vari- 
ous harmonics. 



The construction and initial testing of 
the synthetic system has been completed 
to the point where it is now feasible to 
use the system and to design new measur- 
ing and processing techniques to take ad- 
vantage of this new radar system. 

SYNTHESIZED PULSE 

To understand synthetic pulse radar and 
how it differs from the more conventional 

^Leckenby, R. J. Electromagnetic 
Ground Radar Methods. Paper in Premining 
Investigations for Hardrock Mines. Pro- 
ceedings: Bureau of Mines Technology 
Transfer Seminar, Denver, CO, Sept. 25, 
1981. BuMines IC 8891, 1982, pp. 36-45. 

^Fowler, J. C, S. D. Hale, and R. T. 
Houck. Coal Mine Hazard Detection Using 
Synthetic Pulse Radar (contract H0292025, 
ENSCO, Inc.). BuMines OFR 79-81, 1981, 
84 pp. 



For waves that are not periodic, but 
have zero displacement beyond a certain 
finite range, a Fourier series cannot be 
used. Instead, a Fourier integral is em- 
ployed. The nonperiodic wave is composed 
of (or can be synthesized by adding) an 
infinite number of wave trains, each hav- 
ing a frequency differing only infinites- 
imally from the next. The process of de- 
termining the components at each frequen- 
cy is done by what is known as the 
Fourier transform, and likewise the wave 
synthesis is done by the inverse Fourier 
transform. 

The above discussion is based on 
continuous waves, but in actual prac- 
tice, waves are measured for finite 
periods. This leads to a whole new 
discussion on discrete Fourier trans- 
forms and a computational method known 
as the Fast Fourier Transform (FFT) 
and the Inverse Fast Fourier Transform 



77 



(IFFT),^ which will be described by sim- 
ply stating that a FFT of a discrete time 
series using prescribed cautions, meth- 
ods, and rules will yield a Fourier 
series of finite components which will be 
a good estimate of the actual Fourier 
transform. 

With this general background of the 
Fourier transform it is possible to see 
the difference, or probably better 
stated, the similarity between a short 
pulse and the synthetic pulse. Figure 1 
is composed of three graphs. The top 

^Bracewell, R. N. The Fourier Trans- 
form and Its Applications. McGraw-Hill, 
1978, pp. 356-381 . 

Brigham, E. O. The Fast Fourier 
Transform. Prentice-Hall, 1974, 252 pp. 



graph is a wave contour of a digitized 
transmitted pulse after traveling through 
about 12 m of coal. A FFT was applied to 
this time domain data to yield the fre- 
quency domain data plotted in the lower 
two graphs. The center graph is the am- 
plitude spectrum, which is the magnitude 
of the sine and cosine components at each 
discrete frequency. The bottom graph is 
the phase spectrum which is the tangen- 
tial angle between the two components. 
The wave was digitized with 1,024 sam- 
ples of equal time intervals, between 
to 500 ns. This yielded a corresponding 
spectrum with a fundamental frequency of 
2 MHz and 512 harmonic frequencies, of 
which only the first 65 are plotted be- 
cause the remaining frequencies are so 
low in amplitude that they appear almost 
zero on the graph. 




50 100 150 200 250 300 350 400 450 500 
TIME, ns 




-180, 



12 25 37 50 62 75 87 100 112 125 
FREQUENCY, MHz 



■'■I|' I'l I'l'''''^'' P'l' l' I'll 



25 



37 50 62 75 87 

FREQUENCY, MHz 



100 112 



FIGURE 1. - Three graphs showing the short 
puise signal in the time domain (top graph) and 
the frequency domain (bottom two graphs). From 
top to bottom the graphs are (1) wove contour, 
(2) amplitude spectrum, and (3) phase spectrum. 



In operation each discrete frequency 
needed to synthesize a pulse is transmit- 
ted one at a time, and the resultant am- 
plitude and phase at each frequency is 
recorded. That is, the acquired data are 
from the frequency domain and can be pre- 
sented in a similar fashion as the lower 
two graphs in figure 1. To ascertain the 
time domain or corresponding pulse data 
an IFFT or equation 1 is used to synthe- 
size a wave, which can be presented in a 
fashion similar to the wave contour of 
the upper graph in figure 1. 

SHORT PULSE VERSUS SYNTHETIC PULSE 

The discussion thus far has served to 
explain how a short pulse and synthetic 
pulse will basically produce the same re- 
sults. The major difference is in the 
hardware and how it controls the energy 
of the transmitted and received signals. 
On a short pulse system the pulse is usu- 
ally generated by a fast discharge of a 
high-voltage potential into an antenna 
and a complex impedance. The potential, 
rate of discharge, antenna, and impedance 
control which frequencies and their am- 
plitudes will be transmitted into the 
ground. This in turn will determine the 
shape of the pulse, the resolution and 
the effective pulse power, the sum of the 
power at each frequency, which in turn 
determines the range of the system. 



78 



Figure 1 is a good example of an erratic 
pulse spectrum that can be ascertained 
from a short pulse system. On the other 
hand the synthetic pulse transmits each 
known frequency at a given power, which 
assures better control over the shape, 
resolution, and power of the transmitted 
signals. It is also easier to obtain 
higher effective pulse power than by us- 
ing a broadband short pulse. That is, a 
small increase in power for each frequen- 
cy, when summed, gives a large increase 
in pulse power. Likewise, pulse power 
may be increased by increasing the number 
of frequencies transmitted. 

THE SYNTHETIC PULSE SYSTEM 

The major components of the synthetic 
pulse system are shown in figure 2. The 
system employs a hetrodyne receiver, 
which requires two synthesizers that are 
phase locked together with a 1-mHz off- 
set. By mixing the transmitted signal 
and the intermediate frequency (IF) of 1 
MHz with the received signal, the detec- 
tion circuits only have to operate at the 



IF frequency and not over the entire band 
of transmitted frequencies. This mixing 
allows for easier shielding of the input 
from direct feedthrough of the transmit- 
ted signal. A fiber optic link is also 
used between the control unit and the 
transmitter to eliminate direct feed- 
through signals. 

The transmission frequencies are in the 
range of 20 to 160 MHz, with a mini- 
mum frequency spacing of 100 kHz, thus 
providing a time window of 10,000 ns. 
The control unit, shown in figure 3, is 
housed in an aluminum case with its own 
batteries. The receiver electronics are 
contained in the receiving antenna. The 
transmitter, figure 4, consists of a 
small box containing batteries and elec- 
tronics with the power amplifier being 
mounted in the transmitting antenna. The 
antennas are loaded wideband dipoles sim- 
ilar to those used in short pulse sys- 
tems. A host computer, which is required 
to perform the IFFT, can access the digi- 
tal tape by either a serial or parallel 
interface. 



Tape 
drive 



Control 

micro- 

processor 



A/D 
conversion 



CONTROL UNIT 



Synthesizer 



Fiber optic 
transmitter 



Offset 
synthesizer 



1 MHz IF 
reference 



nrRANSMITTER 



Fiber optic 



cable 



Fiber optic [J Power 
amplifier 




Sample 
and hold 



Quadrature 
hybrid 



Sample 
and hold 



0" 



90' 



Q mixer 



XMIT+IF 
reference 



1st mixer 



Antenna 



Amplifier 



RECEIVER 



Antenna 



^ 



FIGURE 2. - Block diagram of the synthetic pulse radar system. 



79 




-Q 



-a 

c 
a 



-a 

c 

3 

o 



E 

0) 

>. 






3 
O 

a> 

U 

a 

J3 



80 




81 



Receiver station 




170' 



170' 



Transmitter 

FIGURE 5. - Main travel paths for radar sig- 
nals through a 170-ft coal pillar. 

SYNTHETIC PULSE FIELD TESTS 

The synthetic pulse system was tested 
at four separate sites to determine its 
capabilities of detecting hazards in a 
coal seam. Two tests were in Eastern 
coal mines and two were in Western coal 
mines. Figure 5 shows possible travel 
paths from a transmission test through 
170 ft of coal at Consolidation Coal 
Co.'s Humphrey #7 Mine near to Morgan- 
town, WV. These paths include direct 
transmission, reflection off the side 



CO 
UJ 

> 

LiJ 

a 
us 
tr 



30 

20 
10 







^ — AA/n/vk/ 

100 200 300 400 
TIME, ns 



500 600 



FIGURE 6. - Synthetic pulse radar signals 
through 170 ft of coal. 

rib, and refraction along the back rib. 
Figure 6 shows the synthesized pulses 
as the receiver was moved along the back 
rib. Additional "ghost" signals are 
present in this data, because at the time 
of this test the power amplifier was in 
a separate box from the antenna and ring- 
ing occurred in the connecting coaxial 
cable. This problem was later solved by 
mounting the amplifier directly on the 
antenna. 

The field tests were very successful. 
The synthetic pulse system more than 
doubled the probing distance of previous- 
ly tested short pulse systems. The sys- 
tem was also successfully used in deter- 
mining the in situ electrical properties 
of coal. However, the ultimate range and 
limitations of the system have yet to be 
determined. 



SEISMIC METHODS 



Seismic methods in underground coal 
mines are divided into three categories. 
The first uses relatively high frequen- 
cies for near-field, high-resolution of 
smaller reflection targets. The second 
uses guided elastic waves for delineation 
of major anomalies in the far-field. The 
third, a borehole technique, provides 
calibration velocities for the first two 
techniques with the potential for mapping 
coal thickness from a horizontal hole. 



HIGH-FREQUENCY TECHNIQUE 



Operating frequencies in the range of 
5,000 Hz have potential for resolving 
targets as small as 6-in diam within 20 
to 30 ft of the rib or face. Such reso- 
lution is necessary to locate small voids 
(e.g., well bores) and small faults ahead 
of the face. An advantage of the high- 
frequency technique is that the source 
and receiver can be configured to 



82 



provide a directional beam of energy for 
a relatively narrow field of view. Thus 
the boundary effects of reflections from 
the roof and floor are minimized. 

Figures 7 and 8 illustrate a prototype 
system used to demonstrate the high- 
frequency technique. The system was de- 
veloped under a Bureau of Mines contract 
by the Energy and Minerals Research Co. 
In figure 7, the source and receiver 
transducers are shown mounted side- 
by-side on the rib of a coal pillar. The 
source and receiver are piezoelectric 
transducers tuned to the same (5,000 Hz) 
frequency with a relatively low-Q (broad- 
band) response found appropriate to cou- 
ple signal energy into coal media. Elec- 
tronic circuitry (fig. 8) includes a var- 
iable pulse-width drive circuit for the 
source transducer, and signal condition- 
ing circuits for the receiver output. 

The system was calibrated by transmit- 
ting a pulse through the 18-ft-wide pil- 
lar and measuring the traveltime with a 
receiver mounted on the opposite rib 



(fig. 9). Velocity of sound through the 
coal was determined to be 6,922 ft/s, 
consistent with laboratory measurements 
of coal samples from this seam. With 
this information, and by selectively fil- 
tering the received signal on the same 
rib (fig. 7), an easily discernible echo 
from the opposite rib was observed. 
These tests show that it is possible to 
obtain a reflection from a planar inter- 
face at distances of at least 18 ft. The 
detection circuitry, with the velocity 
calibration, lends itself to numerical 
readout of distance to the reflector. 
This will provide an on-site interpreta- 
tion of the results and allow an un- 
skilled operator to scan a volume of coal 
ahead of the face for potentially hazard- 
ous conditions. 

In the preliminary tests of the high- 
frequency system, transducers were bonded 
to the coal rib with resin grout to 
achieve adequate coupling of energy into 
the coal. Subsequent research has empha- 
sized development of a force-insensitive 
mounting configuration, which will be a 




FIGURE 7. - High-frequency piezoelectric transducers mounted on the rib of a coal pillar. 



83 




FIGURE 8. - Breadboard electronic circuitry for pulse shaping, signal conditioning, and detection 
of reflections. 




FIGURE 9. • Small piezoelectric transducer mounted on the opposite rib for velocity calibration. 



84 



hand-held scanning unit employed at vari- 
ous locations to map anomalous conditions 
ahead of the face. It was found that an 
equivalent amount of transmitted energy 
can be achieved with reduced drive volt- 
age on the transmitter by employing gated 
bursts of single frequency, rather than 
single pulses. With lower power levels 
for energy transmission, the system can 
readily be made intrinsically safe for 
operation in coal mines. 

GUIDED WAVES 

Under certain conditions, a coal seam 
may be a waveguide for long-range propa- 
gation of seismic energy. This requires 
a seam that is bounded by roof and floor 
rock that have a greater density and a 
higher sound velocity than the coal. 
When these criteria are met, as they gen- 
erally are in nature (the coal seam being 
bounded by shale or sandstone), then the 
multiply reflected waves from the roof 
and floor will constructively interfere 
to produce resonant modes of propagation 
that undergo less attenuation than the 
direct body waves. These normal resonant 
modes are dispersive (different frequen- 
cies travel at different speeds) and are 
referred to as Rayleigh- or Love-type 
waves because of their similarity to 
earthquake-generated waves, which travel 
large distances over the earth's surface. 

The dispersive nature of the normal 
modes and the restrictive nature of the 
underground environment complicate the 
application of the technique, but the po- 
tential benefits of far-range detection 
of faults and abandoned workings justify 
research to develop the concept. 

Two procedures are used in underground 
mines: through-transmission and reflec- 
tion surveys. Seismic sources used are 
generally small explosive charges placed 
in drill holes or hammer blows on the 
rib. Because explosives present an obvi- 
ous safety hazard, and because hammer 
blows are not reliably repeatable, the 
Bureau developed, under a contract, 
controlled-source piezoelectric trans- 
ducers (fig. 10) to generate predominant- 
ly compressional or shear wave energy. 



The most desirable type of wave from the 
standpoint of simplicity in interpreta- 
tion is a horizontally polarized shear- 
wave, which will be totally internally 
reflected within the seam for the normal 
modes of the Love type. The shear wave 
source and receiver can be mounted on the 
rib of a coal pillar (fig. 11) to prefer- 
entially excite horizontal particle mo- 
tion and generate the Love-type modes. 

The shear wave source and matching re- 
ceiver were used to demonstrate the 
through-transmission and reflection meth- 
ods in a coal pillar at the Bureau's 
Safety Research Coal Mine, Bruceton, PA 
(fig. 12). Waveforms recorded at the re- 
ceiver directly opposite from the trans- 
mitter having the shortest travel path 
are reproduced in figure 13. The top 




FIGURE lOo - Controlled-source shear wave 
transducers. 




FIGURE IL • Shear wave source mounted on 
the rib of a cool pillar to generate horizontal 
vibration. 



85 





« 37.0' i 


11.0' 


■ 50.0' i 












^Detector 

u^n 1 


82 






Detector 
\ 

N 
\ 
N 
\ 

s 




«s 




"^ 





Source 



20 



Scale, ft 

FIGURE 12. - Plan view of the test pillar at the 
Bruceton Mine with transducer locations and trav- 
el paths for seismic signals. 




FIGURE 13. • Waveforms recorded at the nearest 
receiver location. The shear wave source has been 
rotated 180° to obtain the bottom trace, reversing 
the polarity of the arrival indicated by the arrow. 

trace shows a low-amplitude first arrival 
followed by a large secondary signal, 
which Is Interpreted as the shear wave 
traveling In the coal seam at a speed of 



approximately 2,300 ft/s. Support for 
this interpretation is provided by the 
lower trace, which was obtained by re- 
versing the pulse direction of the source 
transmitter. The shear wave arrival 
clearly shows a 180° phase shift (arrows) 
as expected. 

In figure 14, the waveforms recorded at 
the two receiver locations are compared. 
Here the top trace is an amplified ver- 
sion of the previous data for the nearest 
receiver. The directly transmitted shear 
wave arrival is again indicated, and a 
later very similar arrival with lower am- 
plitude and reversed polarity is clearly 
evident at about 100 ms. Arrival time 
for this event corresponds to reflection 
of the shear wave from the end of the 
coal pillar; a polarity reversal would be 
expected from the negative reflection co- 
efficient at the coal-air interface. The 
bottom trace shows the waveform at the 
far receiver position of figure 12. The 
direct shear wave is obscured in the 
early portion of the trace where some ap- 
parent resonance or ringing appears; how- 
ever, after the amplitudes die off, an 
arrival at about 150 ms corresponds to 




FIGURE 14. - Waveforms recorded at the near 
receiver (top) and far receiver (bottom). 



86 



reflection from the end of the pillar 
over the longer travel path for the sec- 
ond receiver position. It should be not- 
ed in figure 12 that the various travel 
paths are at different angles to the 
direction of particle motion excitation; 
for the far receiver position, propor- 
tionately more compressional wave com- 
ponent of energy would be detected, 
possibly contributing to the higher am- 
plitude early arrival and obscuring the 
direct shear arrival. 

BOREHOLE TESTS 

When explosives or hammer blows are 
used in underground seismic surveys, the 
more complex waveforms generated require 
accurate determination of both the com- 
pressional and shear wave velocities for 
interpretation. A borehole sonic log- 
ging probe was adapted for use in a hori- 
zontal drill hole to investigate veloci- 
ties near the edges of a coal panel. The 
probe is a dual-receiver, two-component 
tool designed to selectively detect par- 
ticle motion parallel or radial to the 
borehole axis (compressional or shear 
waves, respectively). Hydraulic pistons 
clamp the transducers in rigid contact 
with the borehole wall, and the probe can 
be used in either fluid-filled or dry 
holes. Figure 15 illustrates the probe 




Transmitter 



Waveforms recorded 
at near detector 



S'Wave P-wave 
channel channel 



Waveforms recorded 
at far detector 





FIGURE 15. - Diagram of dry hole sonic probe 
and sample waveforms. 



configuration, with typical waveforms 
observed on the compressional and shear 
wave channels at the two receivers. 

The logging probe was positioned at a 
starting depth of 17 ft in a drill hole 
in a coal pillar (fig. 16). Transducers 
were clamped to the borehole wall, 
and full waveform recordings were taken 
on the compressional and shear wave 
channels for both receivers , Rl spaced 
at 4 ft and R2 6 ft from the transmit- 
ter. Pistons were then retracted and 
the process was repeated at 1-ft inter- 
vals toward the rib. Velocity deter- 
minations from these data are plotted 
in figures 17 and 18. Both the compres- 
sional and shear wave velocities exhibit 
low values near the rib, reach a shoulder 
a few feet into the rib, and tend to 
level off toward a constant value with 
depth. 



2 3 15 6 Z a 9 10 I! 12 13 14 



4-in-diam 



Start position 
Tx depth (17ft) 



. End position 
Tx depth (4ft) 



in-- 10' 



FIGURE 16. - Cross section view of horizontal 
drill hole in a coal pillar and locations of veloc- 
ity measurements. 




2,000 

2 



4 6 8 10 12 14 16 

SOURCE-RECEIVER MIDPOINT DEPTH, ft 



FIGURE 17. - Compressional wave velocities 
versus depth of measurement in a coal pillar. 



87 



5,000 



o 3,000 




2 4 6 8 10 12 14 16 

SOURCE-RECEIVER MIDPOINT DEPTH, ft 

FIGURE 18. - Shear wave velocities versus 
depth of measurement in a coal pillar. 

The two lower plots in figures 17 and 
18 represent average velocity deter- 
minations over the 4-ft and 6-ft travel 
paths from transmitter to receiver, and 
include the effects of variation in 
borehole diameter and electronic delays 
in the probe circuitry, producing some- 
what low values. The upper curves rep- 
resent differential determination of 
velocities over the 2-ft travel path 
between the two receivers. They should 
represent more realistic values because 
any constant electronic delays are ac- 
counted for; however, the values are 
more variable because of greater sensi- 
tivity to local anomalies in the borehole 
wall. 



4,000 



3,000 



UJ 



2,000 



1,000- 




4 6 8 10 
DISTANCE, ft 

FIGURE 19. • Traveltime curves for determin- 
ing average velocities within the coal pillar. 



Average velocities for calibration of 
the seismic surveys were established by 
plotting a summation of the differential 
traveltimes versus distance along the 
borehole (fig. 19), neglecting the low- 
er velocity values shallower than the 
shoulder at about 4-ft depth in fig- 
ures 17 and 18. A good correlation is 
achieved with this method using linear 
regression to provide values of shear and 
compressional wave velocities which are 
consistent with previous well-log sonic 
data in vertical exploration holes. 



CONCLUSION 



Four geophysical methods — synthetic 
pulse radar, high-frequency seismic, 
guided waves, and borehole velocity log- 
ging — were investigated by the Bureau to 
devise better in-seam hazard detection 
techniques. It is anticipated that these 
techniques will complement each other and 
will provide a valuable tool to the min- 
ing industry. The synthetic pulse radar 
provides high-resolution reflection capa- 
bilities in the range between 10 ft and 
200 ft. The high frequency seismic meth- 
od, when fully developed, should provide 
near-range, up to 30 ft, high-resolution 
reflection detection capabilities, and 



the lower resolution, long-range guided- 
wave method using the control sources 
will provide detection in the hundreds of 
feet. The borehole sonic probe will be 
useful for determining the seismic veloc- 
ities needed for both the high-frequency 
seismic and guided-wave methods. 

Research is continuing on all four 
methods. This paper demonstrates the po- 
tential of the methods. Further re- 
search, development, testing, and actual 
use are required to reach the expectation 
that is expected for each method. 



88 



BIBLIOGRAPHY 

Snodgrass, J. J. A New Sonic Velocity- Suhler, S. A., T. E. Owen, B. M. Duff, 
Logging Technique and Results in Near- and R. J. Spiegel. Geophysical Hazard 
Surface Sediments of Northeastern New Detection From the Working Face (contract 
Mexico. BuMines TPR 117, 1982, 24 pp. H0272027, Southwest Res. Inst.). BuMines 

OFR 69-83, 1981, 176 pp. 



SATELLITE IMAGERY AS AN AID TO MINE HAZARD DETECTION 
By Robert A. Speirer'' and Stanton H. Moll^ 



89 



ABSTRACT 



The Bureau of Mines is involved in 
ongoing research to develop potential 
hazard evaluation maps for mine areas . 
These maps will be generated using 
computer-aided methods to analyze Land- 
sat imagery and multivariate data sets. 
A means of image processing whereby 
lineament information is enhanced and 



extraneous information suppressed has 
been devised. Digital processing is 
particularly appropriate for picking 
lineaments from Landsat data because it 
is faster, less biased, more repeatable 
and, ultimately, less costly than manual 
interpretation. 



INTRODUCTION 



The geologic environment in mining ar- 
eas directly influences the safety of 
mine workers. Where possible hazards 
exist, it is essential that their nature 
and location be determined before mining 
into them. The basic mine plan can then 
be modified at an early date for safety 
and economy. 

Geologic features in coal mines , such 
as faults or sand channels, cause zones 
of weakness in the roof due to fracturing 
or differential compaction. These fea- 
tures usually require that roof bolt 
plans in their vicinity be modified. 
Roof or rib falls in underground coal 
mines and slides in open pits are related 
to these geologic features and still 
account for a large number of fatali- 
ties. Although fatalities were reduced 
since the enactment of the Federal Coal 
Mine Health and Safety Act of 1969, roof 
falls still cause some 40 pet of all 



underground coal mine fatalities and dis- 
abling injuries. 

Detection of possible zones of weakness 
using data derived from Landsat satel- 
lites has been demonstrated by many re- 
searchers. Since the zones of weakness 
that may cause mine hazards may be re- 
flections of discontinuities in the 
earth's crust, surface expressions of 
such discontinuities should be, and of- 
ten are, visible on Landsat images to 
the trained interpreter. Much effort 
has been devoted to mapping lineaments 
(linear features often representing dis- 
continuities) and demonstrating corres- 
pondences between lineaments and natu- 
ral features (i.~2^)»^ Additional effort 
has been applied to the specific task of 
delineating lineaments in mine areas (2) , 
The Mine Safety and Health Administration 
(MSHA) also conducts a program to provide 
technical support for lineament analysis 
to mine operators. 



MANUAL LINEAMENT ANALYSIS 



Current technology consists of a 
trained operator using visual techniques 
to plot linear features onto a base map 
from Landsat images. Rinkenberger (j4) 
used this method to demonstrate the cor- 
relation of lineaments with known faults 
and fractures. Generally, such methods, 
including those used by MSHA, involve the 

'Geologist, Denver Research Center, Bu- 
reau of Mines, Denver, CO. 



use of analog equipment for edge enhance- 
ment and density slicing (mapping ranges 
of brightness to a single color or 
brightness) of subsets of standard Land- 
sat scenes . The visually enhanced images 
are manually interpreted to obtain the 

^Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this paper. 



90 



lineament map, which is then registered 
and overlaid onto a base map. 

An example of a lineament map over a 
mine area is shown in figure 1. The 
lineaments have been manually interpreted 
from a Landsat scene. When overlaid on a 
mine outline, a high correlation is seen 
between the lineaments and known roof 
fall occurrences (small dots). With the 
aid of a lineament map , the mine operator 
has some warning of zones of potential 
hazards and can use the map as a guide in 
anticipating hazardous ground conditions 
(large dots). Verification of these haz- 
ards may lead to installing additional 
roof support or to modifying the mine 
plan to avoid the hazards. 



Although this method has proven via- 
ble, and is generally accepted by mine 
planners, it has a number of substantial, 
interrelated drawbacks. These include: 

1. Repeatability . Repeated interpre- 
tations of the same image are rarely, if 
ever, identical. Two interpreters will 
seldom agree about the length, azimuth, 
or number of lineaments . Even the same 
trained interpreter rarely obtains the 
same results on subsequent trials. 

2. Bias . Each interpreter brings his 
or her own biases to the task. Fur- 
thermore, since these are an unconscious, 
and hence, of unknown quantity, they are 
extremely difficult to control and 




MINE 'a' 



FIGURE 1. - Lineament map registered to a mine map, generated by the manual method. The areas 
where greater roof instability may be expected are shown in large dots. 



91 



compensate for. For example, an inter- 
preter with some previous knowledge of 
the structure of an area will probably 
color the interpretation with those pre- 
conceptions. One interpreter may prove 
adept at locating "local" or short fea- 
tures, whereas another may, for example, 
preferentially locate northeast-trending 
features over north-trending lineaments. 
Again, the individual biases may change 
from trial to trial. 

3. Experience . The interpreter must 
be trained in the "art" of lineament 
analysis. The more experience an in- 
terpreter has , the more he or she 
will be able to control the problems of 



repeatability and bias. At the same 

time, however, the cost of analysis 

will increase with the increase in 
experience. 

4. Speed and cost . Although a Lands at 
scene can be analyzed by a trained inter- 
preter using just an image and a pencil, 
the process is slow and tedious. The 
cost is low because it consists of only 
an interpreter's wages and the cost of 
the image. Enhancement devices, at added 
cost, can be used to speed up the analy- 
sis. Such devices do not, however, re- 
lieve the tedium of the manual interpre- 
tation process nor eliminate the other 
drawbacks . 



COMPUTER-AIDED LINEAMENT ANALYSIS 



Because of the success of the manual 
lineament analysis technique, the Bureau 
is devising a program to improve the 
technique with computer processing. Com- 
puter processing and advanced analytical 
techniques can reduce or eliminate the 
drawbacks of the manual method while im- 
proving throughput and making the tech- 
nology more widely accessible. 

The advantages of using digital meth- 
ods for lineament analysis are several. 
These include: 

1. Suitability . Landsat data is dis- 
crete digital data, and hence lends it- 
self readily to digital processing. 
Images are available from EROS Data 
Center, Sioux Falls, SD, that have al- 
ready been enhanced by digital means to 
Improve image quality. Scenes can also 
be acquired in digital form as Computer 
Compatible Tapes (CCT's) and further 
processed to enhance desirable features 
such as lineaments. 

2. Repeatability . If identical param- 
eters are entered into a program, the 
computer will repeatedly find an identi- 
cal crop of lineaments from the same 
s cene . 

3. Controlled bias . The computer can 
be directed to find features of a given 
size or shape and can be expected to find 



all features for which it has been di- 
rected to search. Different processing 
methods can enhance or suppress certain 
features. These capabilities cannot be 
expected from a human interpreter. Most 
importantly, the biases of the computer 
can be known, whereas those of a human 
interpreter cannot. 

4. Flexibility . Because computers are 
more easily "retrained" than human inter- 
preters, they can be directed to perform 
many different tasks. 

5. Speed and cost . Although initial 
costs for the computer method are like- 
ly to be greater than for the manual 
technique, this disadvantage is soon 
overcome by greater throughput. Increased 
reliability and repeatability, and less 
need for a trained Interpreter. Soft- 
ware development costs may be high for a 
single application, but they can usually 
be amortized and shared among several 
uses. Less time is needed to train an 
operator for the image processing system 
than to train an interpreter for the man- 
ual technique. 

An earlier study, now complete, was 
made to try to devise a computer tech- 
nique to detect lineaments in Landsat 
scenes. After transferring the linea- 
ments to a base map , they were field 
checked to verify the correspondence 



92 



between the plotted lineaments and sur- 
face geology. The technique proved suc- 
cessful in discovering some features in 
the mining areas where it was tested. It 
was apparent, however, that further re- 
finement was needed to learn how to dis- 
criminate between manmade features such 
as roads, canals, fields, vapor trail 



shadows, etc., and natural phenomena such 
as faults, fractures, joints, and paleo- 
channels. Furthermore, the technique 
proved to be "blind" to lineaments in 
certain orientations, and also to other 
nonlinear features that may be important, 
such as circular or serpentine patterns 
(e.g. calderas and thrust faults). 



AUTOLINER PROJECT 



The Bureau has recently concluded an 
interagency project with the U.S. Geo- 
logical Survey to develop a method of 
automatic lineament mapping O ) . The 
Autoliner project, although not capable 
of generating a lineament map per se, did 
result in a method of processing the 
Landsat scenes to enhance those natural 
features of interest to the image analyst 
while suppressing extraneous information. 
The resultant image is far superior to a 
standard image for lineament analysis. 

AUTOLINER METHOD 

Basically, the autoliner works as 
follows: 

A Landsat satellite records the reflec- 
tance of an area on the ground in each of 
four spectral bands. The smallest area 
that can be resolved, called a pixel 
(picture element), measures about 75 m 
by 75 m. Each pixel, in each band, is 
sent back to earth as a value between 0^ 
(for no reflectance) to M^ (for total 
reflectance) . 

On earth the data is cleaned up (pre- 
processed) to correct for such aberra- 
tions as differences in detector calibra- 
tion, atmospheric haze, data transmission 
errors, and geometric distortion. When 
this procedure is complete, the data 
still comprises four bands ranging in 
value from 0^ to W^ each, but now is geo- 
graphically correct. An image made at 
this point would be brighter and easier 
to interpret than an image made before 
preprocessing. These images are marketed 
by EROS Data Center, and are the images 
usually used in manual interpretations. 



A great deal more processing can be 
done, however, to enhance specific fea- 
tures. Since the pixel values (called 
density numbers, or DN) are directly pro- 
portional to reflectivity, the difference 
between pixel values is their contrast. 
Linear features will usually have a mod- 
erate contrast, which will be due to veg- 
etation or elevation differences across 
the feature. Other features, such as 
snow-field boundaries, will have extreme 
contrast, whereas very low contrast is 
usually minor in content. Consequently, 
a window can be specified. Outside this 
window data can be ignored (or set to no 
contrast), whereas contrasts inside the 
window can be magnified to further in- 
crease the contrast. 

There are many ways of establishing the 
contrast of a pixel with its surround- 
ings . A pixel may be compared with a 
single immediate neighbor or with any 
number of surrounding pixels. A small 
area will more closely reflect the value 
of its central pixel than will a large 
area, hence the more distant pixels 
should be more suppressed than the proxi- 
mal pixels. These methods of establish- 
ing and enhancing contrast were studied 
in the Autoliner project. 

The results of the optimal method for 
enhancing lineament information is shown 
in figure 2. Called a "thematic linear 
feature map," the values within the se- 
lected window are shown in black, and all 
contrast values outside the window are in 
white. Selected lineaments are shown by 
pairs of arrows. 



93 




FIGURE 2. - Thematic (binary) lineament feature map of a Denver Landsat 3 image, 
produced using the modified-modified gradient method. Arrows show lineaments. 



AUTOLINER CONCLUSIONS 

The Autoliner project resulted in prom- 
ising techniques of image processing and 
enhancement for lineament analysis. 
These techniques permit the highlighting 
of data that contain lineament informa- 
tion, while suppressing information ex- 
traneous to the interpretation. However, 
the procedures outlined so far must still 
be interpreted by a trained lineament 
analyst. That is, they are still subject 
to visual analysis for lineament picking, 
albeit with a much improved product. 

The image processing software used in 
the Autoliner project will shortly become 



available to the public. MIPS (Minicom- 
puter Image Processing System) was de- 
signed to operate on a DEC PDP-11 mini- 
computer, model 23 or higher, running the 
RSX-llM operating system. Color graphics 
are provided by a Grinnel model GMR 27 or 
270 image processing display system, with 
hardcopy output via color Optronics or 
high-resolution Dunn camera.-^ 



^Reference to a specific brand, equip- 
ment, or trade name in this report is 
made to facilitate understanding and does 
not imply endorsement by the Bureau of 
Mines. 



94 



Most of the processing routines in MIPS 
are written in DEC FORTRAN IV PLUS, and 
hence should be readily transportable to 
other computers. However, since image 
processing deals with such large data 
sets, the decision was made to optimize 
I/O (data input and output operations) 
and data transformation routines by 



writing them in machine-specific assembly 
language, and using low-level operating 
system routines. Consequently, translat- 
ing these routines to another computer 
will involve some investment of time and 
probably greater program execution time 
as well. 



FUTURE RESEARCH 



The primary objective of Bureau re- 
search with the MIPS system is to develop 
a potential hazard map for use by mine 
planners. A similar product acquired by 
the manual method was shown in figure 1. 
Using the MIPS system, it is hoped that 
this product can be improved in two ways : 
(1) by incorporating the "Autoliner" 
methodology and its derivatives, and (2) 
by integrating other data sets into the 
hazard analysis. 



Figure 3 shows a prototype of a hazard 
map generated using digital means. Al- 
though numbers have not , at this point , 
been assigned to the contours, they could 
represent a variety of quantities, such 
as degree of hazard or length of roof 
bolts to be installed. Assessing the 
parameters to incorporate into the plan, 
and their respective weights, is one of 
the objectives of the Bureau's research. 











Potential 
hazard 

□ High 

□ Med 

□ low 



FIGURE 3. - Preliminary potential hazard map of the some area as shown in figure 1, to be gen- 
erated by integration of Autoliner results and geologic, geophysical, geochemical, and miscellane- 
ous data sets. Contours shown represent arbitrary units and values. 



95 



For example, lineament analysis alone 
cannot provide data about the degree of 
weakness each lineament represents. For 
this information, ancillary and comple- 
mentary data sets must be used. Part of 
the Bureau's research will be to assess 
the applicability of these data sets. 



Data sets which are anticipated as being 
useful include radar, magnetics, gravity, 
digital elevation data, geologic mine 
maps, previous roof -fall data, drilling 
data, methane and helium concentrations, 
and other geophysical, geochemical, geo- 
logical, and miscellaneous data. 



CONCLUSIONS 



Computer applications are becoming com- 
monplace in mine planning, and we expect 
that they will soon be common in daily 
operations as well. Furthermore, given 
the massive amounts of mine-related data 
available, the only feasible means of 
assimilating it is by computer. Our re- 
search was undertaken to assist mine 
planners and operators in promoting safe 
and economic operating conditions. 



Since much of the current work in pro- 
viding lineament information to mine 
planners is done by MSHA, the Bureau in- 
tends to work closely with MSHA in devel- 
oping the methodology. It is hoped that 
eventually MSHA will have image process- 
ing systems in their field offices where 
a semi-automatic analysis can be generat- 
ed locally for mine operators. 



REFERENCES 



1. Lillesand, T. M. , and R. W. Kiefer. 
Remote Sensing and Image Interpretation. 
Wiley, 1979, 597 pp. 

2. Short, N. M. The Landsat Tutorial 
Handbook. NASA Ref. Publ. 1078, 1982, 
553 pp. 

3. Burdick, R. G. , and R. A. Speirer. 
Development of a Method To Detect Geo- 
logic Faults and Other Linear Features 
From LANDSAT Images. BuMines RI 8413, 
1980, 74 pp. 



4. Rinkenberger , R. K. Implementing 
Remote Sensing Techniques for Evaluating 
Mine Ground Stability. Mining Enforce- 
ment and Safety Administration (now Mine 
Safety and Health Administration), Inf. 
Rep. 1057, 1977, 34 pp. 

5. Chavez, P. S., Jr. Autoliner Pro- 
ject. BuMines Interagency Agreement 
J0215036; for inf., contact R. A. 
Speirer, TPO, Denver Research Center. 



96 



MICROSEISMIC TECHNIQUES APPLIED TO FAILURE WARNING IN MINES 
By Fred W. Leighton^ 

ABSTRACT 



Miners have long known that rock noise, 
or the popping and cracking of the rock 
commonly heard during mining, can be in- 
dicative of the stability of the mine 
structure. For many years, miners have 
"listened" to the rock talk and many 
times have interpreted changes in rock 
noise activity to be a warning of fail- 
ure and have retreated from the failure 
area. Microseismics , or the study of 
rock noises, was begun in the early 
I940's, partly because of this historical 
fact. 

Microseismics uses geophysical equip- 
ment to detect and analyze rock noises on 
both the audible and subaudible level. 
Thus, these systems are much more sensi- 
tive than the human ear and "hear" even 
more of the rock "talk" than do miners. 



Research has shown that microseismics 
can be used to precisely locate those 
portions of a working area that are gen- 
erating rock noise, and that the rock 
noise release rate information from each 
area can be used to analyze its stabil- 
ity. In some instances, rock noise data 
have been analyzed to provide warning of 
imminent structural failure, and person- 
nel have been removed from or prohibited 
from entering a failure area. This paper 
presents a brief history of how micro- 
seismics evolved, explains why the tech- 
nique works, and describes the basic 
equipment used. Past results in both 
coal and metal and nonmetal mining sys- 
tems are shown, and recent results con- 
cerning the occurrence of a failure in a 
coal mine advancing longwall section are 
presented. 



INTRODUCTION 



When a rock mass is subjected to chang- 
ing stress conditions, such as those 
caused by mining, small-scale adjustments 
occur within the rock that release seis- 
mic energy. This energy, when in the 
audible range, is called rock noise. 
Those areas in which stress changes occur 
are also the areas of the structure most 
likely to fail. Individual rock noises 
can be detected and analyzed to determine 
their precise location relative to the 
mine structure. Over a period of time, 
plots of these data provide a pictorial 
representation of where stress changes 
are occurring, because rock noise activ- 
ity tends to concentrate in those areas 
of the structure most actively adjusting 
to the changing stresses. Since these 
areas also represent those areas most 
likely to fail, potential failure areas 
can be pinpointed and mapped relative to 

^Supervisory mining engineer, Denver 
Research Center, Bureau of Mines, Denver, 
CO. 



the structure. Rock noise rates, or the 
number of rock noises occurring per unit 
of time, also tend to vary dramatically 
prior to failure. Thus, the ability to 
locate the source of individual rock 
noises provides a means of determining 
where failure may occur, and rate count- 
ing within each area offers a means of 
assessing when failure may occur. This 
information, properly treated, can be 
used as an aid to help avoid, control, or 
provide warning of impending failure. 
Successful applications of this technique 
have provided recent impetus to the ef- 
fort of developing microseismics into a 
practical, reliable, and economically 
feasible tool. 

The phenomenon of naturally occurring 
rock noise was discovered in 1938 by 
Obert (J_),2 who was measuring seismic 

^Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this paper. 



97 



velocities in mine pillars. Seismic 
energy other than that which he was gen- 
erating continually registered on the 
recording equipment. Further study by 
Obert and Duvall showed that the extrane- 
ous seismic energy being recorded was 
from rock noises that were being gener- 
ated naturally within the rock in highly 
stressed areas. Pursuing this interest- 
ing phenomenon, both in the laboratory 
and in the field, they documented the 
dramatic change in rock noise rate prior 
to failure (^~3^) • Carrying this work 
further, they established the fact that 
in many instances, rock burst failures 
could be predicted by monitoring and 
"listening" to the rock noise activity in 
rock-burst-prone areas (2^). 

These early efforts clearly showed that 
microseismics had great potential as a 
tool for measuring or estimating the sta- 
bility of mine structures. Extended 
testing, however, showed that sometimes 
rock bursts occurred with no apparent 
microseismic warning, and at other times, 
sharp increases in microseismic data were 
not accompanied by failure. Also, be- 
cause precise location of individual rock 
noises was not at that time possible, one 
never knew where the failure was going 
to occur, only that failure nearby the 



observation point was likely. Thus, 

while the technique clearly offered 

promise, it was not at that time consid- 
ered practical. 

In the mid-1960' s, the Bureau of Mines 
began a new research effort to improve 
the microseismic technique. Major im- 
provements were judged possible, in great 
part due to the availability of new and 
vastly improved electronic and system 
components which had resulted from in- 
strumentation developed during the space 
program. Thus, in 1967, development of 
a multichannel, broadband, microseismic 
system began (^) • Application of this 
system in rock-burst-prone mines showed 
it to work well and provided the incen- 
tive to develop methods whereby the 
source location of individual rock noises 
could be easily and directly calculated 
0-_7) . The improved system and the abil- 
ity to locate rock noises showed through 
application that the microseismic tech- 
nique offered new and increased potential 
as a useful engineering tool for the min- 
ing industry (8^) . The following sections 
briefly describe the basis for microseis- 
mics and current microseismic systems, 
and offer examples of how the technique 
has been and is being applied to mining 
problems and failure warning research. 



HOW MICROSEISMICS WORK 



As has been stated, mining results in 
ever-changing load and stress conditions 
in the ore body and in the mine structure 
support system. These changes are accom- 
panied by rock noise, or the release of 
low-level seismic energy, which can be 
detected by geophones placed throughout 
the mine structure. Each individual rock 
noise can be accurately located relative 
to the mine structure, thus delineating 
those areas most actively adjusting to 
mining. This process is simple and 
straightforward. 

An example of this phenomenon is shown 
in figure 1 , which depicts a plan view of 
a rock noise occurring in the barrier 
pillar of a room-and-pillar retreat sec- 
tion in a coal mine. The seismic energy 
released by the rock noise travels out- 
ward from the source of the rock noise 
in all directions at some measurable 



velocity (not necessarily the same in all 
directions), thus arriving at different 
geophone locations at different times, as 
shown in the example seismic record in 
the lower portion of the figure. Knowing 
the coordinates of each geophone , the 
velocity at which the seismic energy 
travels, and the time at which each geo- 
phone in the array sensed the arrival of 
the seismic energy allows one to calcu- 
late the coordinates of the rock noise 
source. Each source is then plotted on 
the mine structure map , and those areas 
most actively adjusting to the new stress 
regime are delineated by the areas in 
which rock noise activity is most dense. 
In this manner, potential failure areas 
are delineated and pinpointed relative to 
the mine structure. 

Figure 2 shows such a rock noise 
concentration in the two orthogonal 



98 



Rock noise source 




Seismic wave 



Geophone 



A Example geopiione array 



2 

3 

4 
5 

6 



/A/^/VA^.A ^. 



-A^A/^AAr- 



t. 



-V\A/V- — -' — 



t. 



-V\A'*^ 



- — A/\/v^ 

TIME — ► 

B Example record of rock noise 

FIGURE 1. - Rock noise. A, Plan view of a 
geophone array and rock noise in a coal mine 
room-and-pillar section. B, Example of arrival 
time information used to calculate rock noise 
source location. 



. (1 ' ■• •■ ■• 


p 


'• \ ' \ ''wi •' 


t 




1 

1 


• • 





* 



FIGURE 2. - Elevational views of rock noise 
concentration in a rock-burst-prone pillar. 



elevation views of a rock -burst-prone 
pillar (a rock burst is a sudden massive 
and sometimes catastrophic failure) . 
This concentration happened over a 
several-hour period and precisely located 
the area of a rock burst that occurred at 
the beginning of the day shift in the 
area. This example shows the importance 
of being able to locate where rock noises 
are being generated and demonstrates the 
ability of the technique to accurately 
delineate problem areas. 

Another feature of rock noise activity 
is that the rate at which it occurs tends 
to fluctuate dramatically in the area 
prior to failure, which in many instances 
provides a warning of the failure. Fig- 
ure 3 shows the rock noise rate, or num- 
ber of rock noises recorded per unit of 
time, for the location data shown in 
figure 2. Note the dramatic increase in 
the rock noise rate prior to the rock 
burst. 



120 



90- 



z 

i 60 

HI 

(A 
O 

Z 

o 
o 

cc 

30 



T 



Small burst 




J L 



_L 



2 4 6 8 10 12 14 16 

TIME, days » 

FIGURE 3. - Rock noise rate prior to a rock burst. 



99 



These two analyses procedures — i.e., 
event location and rock noise rate 
counting — in combination represent the 
way in which microseismic techniques are 
used in current practice, when using mul- 
tichannel microseismic systems. As will 



be shown, worthwhile applications of 
single-channel systems to monitor local- 
ized areas of interest are also possible, 
using the rate counting method of analy- 
sis without regard to the location of in- 
dividual events. 



MICROSEISMIC SYSTEMS 



Microseismic systems may vary somewhat 
according to the dictates of their appli- 
cation, but essentially all systems in- 
clude the same fundamental components. A 
detailed examination of microseismic sys- 
tems may be found elsewhere (9-11) , and 
only an overview is presented here. 

Figure 4 shows a block diagram of a 
basic multichannel microseismic system. 
Each channel consists of a sensor, a pre- 
amplifier, the data transmission cable, 
and sometimes a postamplif ier to condi- 
tion the signal for final recording. The 
sensor may be either an accelerometer or 
a velocity gauge, depending on the appli- 
cation. Each channel is connected to a 
multichannel, high-speed magnetic tape 
recorder and/or an automatic monitoring 
system that electronically measures the 
necessary information from the seismic 
signals. The data measured by this sys- 
tem are fed into a minicomputer or micro- 
processor , where the data are analyzed to 



Sensor 



Preomp 



High 

gain 

amplifier 





FM to 


pe 


















recorder 


























L- 


3sciMograph 






















RBM 




Computer 




Plotter 














1 








Printer 





FIGURE 4. - Block diagram of a multichannel 
microseismic system. 



compute the coordinates of the source of 
each rock noise, which is then plotted on 
a map of the mine. 

Simpler versions of microseismic sys- 
tems are available for application where 
"listening" is carried out using only one 
channel of equipment. Systems such as 
these are used in certain specific in- 
stances such as in roof fall warning mon- 
itoring, where the area of interest is 
well defined and limited in size. Exam- 
ples of how these systems can be used are 
discussed in "Field Applications." 

Microseismic systems are now available 
commercially, either as a total system or 
by purchasing and assembling individual 
components . While these systems are not 
difficult to use, they do require full- 
time attention and maintenance and are 
not to be considered a "turnkey" opera- 
tion requiring little or no ongoing 
commitment of personnel and capital ex- 
penditure. Sensors are sometimes lost 
during mining and must be replaced, and 
data transmission lines are often severed 
or damaged, requiring attention. The 
data recorded require daily, preferably 
continuous, analysis and interpretation 
to be of maximum value. Thus, a micro- 
seismic system should be considered a 
tool to be applied to stability problems 
and requires the same dedication in terms 
of attention and maintenance as do other 
tools used in the mining routine. Sev- 
eral applications of these systems have 
shown this effort to be worthwhile and 
contributory to a safer and more pro- 
ductive mine. 



100 



FIELD APPLICATIONS 



Microseismic systems have been in- 
stalled and are in use at a number of 
locations throughout the world. Applica- 
tions vary from those using single- 
channel "listening" systems and rate 
counting methods to those incorporating 
24 or more channel systems and the com- 
bination of source location and rate 
counting methods. The following is 
therefore broken into two discussions; 
i.e., those applications using single 
channel systems, and those using multi- 
channel systems. 

SINGLE-CHANNEL APPLICATIONS 



In terms of failure warning, highly 
promising research has recently been done 
using equipment sensitive to the fre- 
quency range of from 40 to 100 Hz (12- 
13). A major advantage of this system 
over the systems sensitive to the lower 
frequency ranges is that manmade noise, 
such as that due to mining, is comprised 
mostly of frequencies lower than 400 kHz, 
hence the data recorded are mostly, if 
not entirely, made up of noises naturally 
occurring in the mine structure. The 
system can thus monitor right in the 
working area, where there is the highest 
risk for personnel. 



Single-channel microseismic systems are 
designed for monitoring well-defined 
problem areas of limited extent. These 
systems detect rock noise activity in 
several frequency ranges, the most common 
being in the 100- to 5,000-Hz range or 
the 40- to 100-Hz range. An advantage of 
single-channel units is that they are 
portable and may be used for periodic 
sampling of many areas within the mine, 
or for continuous monitoring at the work- 
ing area. The disadvantage is that the 
source of the rock noise is never known, 
hence the precise location of a potential 
failure cannot be delineated. As will be 
shown, this disadvantage does not pre- 
clude beneficial use of single -channel 
technology in many instances. 



The system as discussed below was con- 
structed specifically to provide warning 
of impending roof falls. It was designed 
to provide coverage of about a 50- or 
100-ft radius from the sensor, so that it 
would monitor only the working area mini- 
mizing the importance of individual rock 
noise locations. When a warning was 
sounded, one would simply evacuate the 
area immediately surrounding the sensor. 
In practice, this system has been found 
to perform well with large-scale roof 
falls . Figure 5 shows the commercial 
prototype of this device. The system is 
a stand-alone design, providing continu- 
ous monitoring, automatic data analysis, 
and warning of failure independent of 
human input . 



Single-channel data may be recorded and 
used in a variety of ways, ranging from 
counting noises heard through a set of 
earphones to permanently recording and 
analyzing data using sophisticated elec- 
tronics. In the former instance, the 
success of the application relies heavily 
on the dedication and skill of the ob- 
server. In the latter case, the system 
can perform its function essentially 
unattended. Simple systems using head- 
phones and an observer may cost in the 
$1,000 range, while more sophisticated 
systems may cost $10,000. The choice of 
which system to use and consequently how 
much money to invest is dictated by the 
application. 



This system differs from previous 
single-channel monitoring systems in that 
it measures not only the number of events 
that occur within its range, as do stan- 
dard systems, but also estimates the 
total amount of energy released by those 
events . The event count and energy value 
are accumulated for 60-s intervals, and 
then the energy value is divided by the 
total event count to obtain the energy- 
event ratio. This ratio calculation is 
unique to this instrument. In applica- 
tion, the energy-event ratio behaves 
anomalously before failure, and the anom- 
aly is sufficiently large to be easily 
detectable and to be used as a failure 
warning. 



101 




FIGURE 5. - Components of a high-frequency, single sensor microseismic system. 



Figure 6 shows the behavior of the 
energy-event ratio before a large scale 
roof fall (fig. 6A) and before a coal and 
gas outburst, a coal mine failure similar 
to a rock burst (fig. 6B ) . In both in- 
stances, the anomaly is large and occurs 
sufficiently prior to the actual failure 
to ensure removal of personnel from the 
failure area. 

This system has shown much success in 
providing warning of large-scale roof 
falls. Its efficacy as a warning device 
for small-scale roof falls, coal and gas 
outbursts, and other types of failures 
has not yet been determined, but it has 
been shown to work on some occasions. 



These areas of application are the 
focus for current research efforts by 
the Bureau of Mines with the goal of 
establishing the procedures to ensure 
proper use of the device in the field 
and its reliability in different 
applications. 

Single-channel applications of this 
device and others have been historically 
proven to be of value and to be capable 
of providing meaningful information about 
the mine structure in specific areas of 
limited size. To carry out similar stud- 
ies over a larger area, even up to the 
size of the entire mine, multichannel 
systems are necessary. 



102 




4:00 6:00 



z 

LU 

> 
Ui 

>^ 

oc 

UJ 

z 

UI 



8:00 10:00 
TIME, h 



12:00 



8,834 
7,848 
6,867 
5,886 
4,905 
3,924 
2,943 
1,962 
981 





1 

Event 



8:30 



9:00 



9:30 
TIME, h 



10:00 10:30 



FIGURE 6. - Energy-event ratio prior to a 
roof fall {A) and an outburst (6). 

MULTICHANNEL APPLICATIONS 

Multichannel microseismic systems were 
developed by the Bureau in the late 
I960' 8, and research is continuing to 
demonstrate their usefulness in solving 
specific mine failure problems and in 
improving total systems reliability. 
Meanwhile, several applications of the 
technology have been made at its present 
level of development. 

Early examples of systems application 
can be seen in figures 7 and 8, which are 
Bureau of Mines research results (12, 
14) . The important difference between 
single-channel applications and multi- 
channel monitoring is that the location 
of each individual rock noise can be cal- 
culated and plotted on mine maps with the 
latter system. Thus, a larger area of 
the mine, even the entire mine if neces- 
sary, can be monitored on a continuous 
basis, and potential failure areas can be 



accurately pinpointed within the mine 
structure. This capability sometimes 
allows for mine support to be modified to 
avoid failure, for the application of 
destressing techniques to control fail- 
ure, or for the removal of personnel from 
the area prior to failure. 

Figure 7 shows the rock noise activity 
plots for a 5-day period before a rock 
burst in a stope pillar in a metal mine. 
As can be seen, this procedure precisely 
defined the location of the eventual 
failure. Figure 7F^ shows the rock noise 
rate which describes how rock noises in 
the potential failure area occurred as a 
function of time. As in single-channel 
applications, note the dramatic increase 
in rock noise rate prior to the rock 
burst. 

A similar example, this time comprised 
of data from a Bureau study in a coal 
mine, shows the rock noise data recorded 
in conjunction with a coal mine bounce (a 
failure very similar to a rock burst) 
(fig. 8). The rock noise activity has 
been contoured in terms of density, so 
that the inner contours represent the 
area in which the most rock noise is 
occurring. Again, over a period of days, 
notice the dense pattern of rock noises 
that occurred in the eventual failure 
area, precisely defining its location. 
Also, note the similar reaction of the 
rock noise rate from the failure area be- 
fore its failure, as shown in figure 8F^. 

Another recent example of results, from 
the current Bureau study in a longwall 
coal mining system, is shown in figure 9. 
The cumulative rock noise location data 
shown in figure 9A, were recorded during 
the period April 7-14, 1983. A major 
bounce occurred in this section on April 
20, 1983, and caused damage in the tail- 
gate and along the face exactly in the 
area delineated by the rock noise data. 
The rock noise rate also changed during 
this time period, indicating unusual be- 
havior of that face area. 



103 



279 



Ji 



• • 



284 279 



» ^^ • • 

• ■ 



284 



May 20 



279 



' , y v 



279 



May 22 



B 



284 279 



May 21 





/// • 


• 




* 1 i 

• • • 

* • * 




. • .» ^... . .. 




• 
• 


• 



284 



May 23 






284 



o 


lUU 

80 


- 


1 


1 


I 


1 
Burst; 




60 


- 








/- 


Si 


40 


- 








/ - 


0^4 














Oc5 


20 


_ 








/ 


z 


n 






9- 


J — - 


1 




^9 


20 


21 


22 


23 2 


F 








MAY 1979 





F May 24 

FIGURE 7. - Elevational views of rock noise concentrations and rate change prior to a rock burst. 



This data set is incomplete, owing to 
Irregularity in the monitoring procedure 
during this time period, and the pre- 
history and posthistory of this isolated 
data set are unknown. The data cannot 
therefore be said to have offered con- 
clusive evidence regarding failure warn- 
ing. The data are important however, in 
that they (1) provide additional support 
to the hypothesis that bounce areas can 
be delineated well in advance of their 
failure, and (2) indicate the viability 
of these techniques In the mechanized 



longwall mining environment, an important 
contribution in light of the growing num- 
ber of longwalls in the coal mining 
industry. 

The above examples were obtained using 
small geophone arrays comprised of many 
sensors to obtain precise locations of 
rock noises and to provide precise fail- 
ure data of specific areas of interest. 
Similar applications are made in indus- 
try, but the geophones are more widely 
spaced to obtain widespread coverage, and 



104 




April 3-30, 1973 



15 20 
APRIL 



LEGEND 
Event ^ Active mining 

Number ^caved 
3 ' of events 

FIGURE 8. - Plan views of rock noise concentrations and rate change prior to a coal mine bounce. 



105 




B 



11 14 
APRIL 



Headgate 



FIGURE 9„ = Plan view of rock noise concentration and rate change before a coal mine bounce in a 
longwall section. 



accuracy of rock noise locations is only 
valid to about a 40-ft limit. This 
proves sufficient for monitoring several 
stoping areas and provides information 
relavent to the whole of the stope pil- 
lar. This practice, called minewide mon- 
itoring, has been and is presently being 
used with moderate success around the 
world. This type of monitoring provides 
the potential to evacuate a specific 
working area before a burst, as can be 



seen from data such as shown in figure 
10, a case history from the Star Mine, 
Burke, ID (15). It also provides an 
insight into general mine stability and 
an opportunity to watch the development 
of problem areas and to take remedial 
action to control or avoid failure. In 
the Couer d'Alene mining district of 
northern Idaho this type of analysis 
has successfully been used to deter- 
mine when destressing techniques should 



106 



be Initiated to control rock -burst-prone 
pillars (16). CO 



Similar applications have been made in 
other parts of the world for both re- 
search and production. Research efforts 
have increased dramatically around the 
world within the last 5 yr, in both hard 
rock and coal mining. Mining companies 
have installed available systems to moni- 
tor and study specific problems, and in 
some instances are carrying out their own 
research efforts to study new problems. 
These new efforts, on many fronts, will 
have a major impact on the future of 
microseismic techniques in the mining 
industry. 



UJ 

> 

UJ 



00 

3 




18 



FIGURE 10. 
a rock burst. 



12 14 

TIME, h 

Rock noise rate change prior to 



SOME PROBLEMS 



The microseismic technique has devel- 
oped into a potential tool that can pro- 
vide for vastly increased safety to mine 
personnel and can be used as an aid in 
production planning and problems. To 
demonstrate this, clear-cut and highly 
successful applications have been shown; 
however, problems and deficiencies of the 
present technique do exist. 

Reliability is the foremost of the 
problems with microseismic techniques at 
their present level of development. 
Reliability does not mean equipment reli- 
ability, although microseismic systems do 
require constant maintenance and atten- 
tion, but rather tlie fact than not all 
failures are predicted as easily or 
clearly as those presented, and sometimes 
indicated failures may not occur. This 
presents the user with the problem of 
determining how many times he is willing 



to be wrong. Experience has shown that 
the technique has undeniable potential 
and is beneficial in its present form 
when properly used. That same experi- 
ence, however, shows that at times, false 
alarms will be sounded, and even worse, 
sometimes no alarm will be sounded when 
one was necessary. 

The reliability problem cannot be 
attributed solely to improper use of 
microseismic systems or inadequate anal- 
ysis of their data. The solution to the 
problem lies in further research to de- 
velop better and more accurate methods of 
data analysis, and possibly the incorpor- 
ation of supplemental methods to be used 
in conjunction with the application of 
microseismic techniques. While not per- 
fect, the positive aspects of the method 
generally outweigh the potential prob- 
lems to the user. 



CONCLUSIONS 



The microseismic technique has been 
developed a great deal in recent years 
and has been used with moderate success 
to provide warning of failure in several 
different applications in the mining in- 
dustry. Its use is growing and more 
potential users of the technique are 
becoming aware of its capabilities and 
are making efforts to use the technology. 
As mining continues to greater depths and 



stability problems become more intense, 
this use can only be expected to expand, 
and with expanded use, we can expect 
expanded technology and improved methods 
of utilization. 

Current Bureau research, in both hard 
rock and coal mine environments, is aimed 
at establishing better utilization of the 
microseismic technique. Results continue 



107 



to indicate the viability of the tech- warning of failure. The continued ef- 



nique and that its use can contribute 
substantially to both safety and mine 
design. The ability to delineate problem 
areas before failure is becoming better 
established, but the overall reliability 
of current techniques remains undefined, 
particularly with regard to providing 



forts of the Bureau are aimed at estab- 
lishing the data base and experience 
necessary to evaluate present reliability 
problems and to improve the overall 
effectiveness of microseismic techniques 
applied to the problems of the mining 
industry. 



REFERENCES 



1. Obert, L. Measurement of Pressures 
on Rock Pillars in Underground Mines. 
Pt. I. BuMines RI 3444, 1939, 15 pp. 

2. . Use of Subaudible Noises 

for Prediction of Rock Bursts. BuMines 
RI 3555, 1941, 4 pp. 

3. Obert, L. , and W. I. Duvall. The 
Microseismic Method of Predicting Rock 
Failure in Underground Mining. Part II. 
Laboratory Experiments. BuMines RI 3803, 
1945, 14 pp. 

4. Blake, W. , and F. Leighton. Re- 
cent Developments and Applications of the 
Microseismic Method in Deep Mines. Ch. 
23 in Rock Mechanics — Theory and Prac- 
tice. AIME, 1970, pp. 29-443. 

5. Leighton, F. , and W. Blake. Rock 
Noise Source Location Techniques. Bu- 
Mines RI 7432, 1970, 14 pp. 

6. Leighton, F., and W. Duvall. A 
Least Squares Method for Improving the 
Source Location of Rock Noise. BuMines 
RI 7626, 1972, 19 pp. 

7. Redfern, F. R. , and R. D. Munson. 
Acoustic Emission Source Location — A 
Mathematical Analysis. BuMines RI 8692, 
1982, 27 pp. 

8. Blake, W. Microseismic Applica- 
tions for Mining — A Practical Guide (con- 
tract J0215002). BuMines OFR 52-83, 
1982, 208 pp.; NTIS PB 83-180877. 

9. . An Automatic Rock Burst 

Monitor for Mine Use. Paper in Proc. 



Conf, on the Underground Mining Environ- 
ment, Univ. MO, Rolla, MO, 1971. Univ. 
MO— Rolla, 1971, pp. 69-82. 

10. Blake, W. , F. Leighton, and W. 
Duvall. Microseismic Techniques for 
Monitoring the Behavior of Rock Struc- 
tures. BuMines B 665, 1974, 65 pp. 

11. Coughlin, J. P. Software Tech- 
niques in Microseismic Data Acquisition. 
BuMines RI 8961, 1982, 51 pp. 

12. Leighton, F., and B. Steblay. 
Applications of Microseismics in Coal 
Mines. Paper in Proc. 1st Conf. AE/MS 
Activity in Geologic Structures and Mate- 
rials (PA State Univ., June 1975). 
Trans. Tech. Publ., 1977, pp. 205-229. 

13. Steblay, B, Progress in the 
Development of a Microseismic Roof Fall 
Warning System. Paper in Proc. 10th An- 
nual Institute on Coal Mining Health, 
Safety & Research. VA. Polytech. Inst, 
and State Univ., Blacksburg, VA, 1979, 
pp. 177-195. 

14. Leighton, F. A Case History of a 
Major Rock Burst. BuMines RI 8701, 1982, 
14 pp. 

15. Langstaff, J. J. Seismic Detec- 
tion System at the Lucky Friday Mine. 
World Min., Oct. 1974, pp. 58-61. 

16. Blake, W. Rock Burst Research at 
the Galena Mine, Wallace, Idaho. BuMines 
TPR 39, Aug. 1971, 22 pp. 



108 



MECHANICAL AND ULTRASONIC CLOSURE RATE MEASUREMENTS 
By Roger McVey^ 



ABSTRACT 



The Bureau of Mines has constructed two 
intrinsically safe closure rate instru- 
ments that provide the mine operator a 
means for predicting an imminent roof 
fall during pillar robbing. This im- 
proves operator and machine safety and 
prevents delays in digging out equipment. 
One instrument system consists of two 
rugged retrievable extensometers con- 
nected by long electrical cables to a 
digital readout unit for reading closure 
and closure rate. Once a predesignated 
closure rate is reached, the extensometer 
is retrieved by pulling it from the im- 
minent roof fall area by its electrical 
cable. The equipment and mine personnel 
are also pulled back to await the fall, 
which usually occurs within minutes after 



the designated rate is reached. Although 
the unit is primarily designed for re- 
treat mining operations , it can be used 
for any activity requiring measurement of 
displacement or rate of displacement. 
Measurement range is to 6 in with 0.1 
pet accuracy for openings of 4-1/2 to 
12 ft. 

The Bureau is also evaluating a small 
ultrasonic unit to make these measure- 
ments. The new instrument provides unob- 
structing measurements up to 35 ft. The 
ultrasonic transducer can be attached to 
a roof bolt, tossed into an unsupported 
area, or handheld. Total distance and 
rate of change are displayed digitally to 
0.001 ft. 



INTRODUCTION 



Roof control is a major problem in all 
aspects of underground mining, especially 
in room-and-pillar retreat operations. 
Room-and-pillar retreat mining is begun 
by developing a state of multiple en- 
tries. The coal pillars between the 
entries and crosscuts are then extracted 
in a retreat sequence and the roof 
allowed to cave in. The key is to mine 
as much of the pillar as possible, then 
remove both personnel and equipment be- 
fore the final portion of the roof 
collapses. 

An extensive study had been made ear- 
lier at the Southern Utah Fuel Company 
(SUFCO) No. 1 Mine by Hamid Maleki, Colo- 
rado School of Mines, and Doug Johnson, 

^Supervisory electronic technician, 
Spokane Research Center, Bureau of Mines, 
Spokane , WA . 



of SUFCO, in determining the critical 
roof-to-floor closure rate for predicting 
a roof -caving during retreat mining. 
They timed the roof-to-floor clo- 
sure change to determine the critical 
rate of closure. The critical rate was 
determined to be 0.2 in/min. Roof -caving 
prediction based on this value was so 
successful that the mine suggested the 
Bureau develop an automatic closure-rate 
instrument. 

Subsequently, the Bureau designed and 
built two types, mechanical and ultra- 
sonic, of automatic closure-rate instru- 
ments that would digitally record both 
the rate of closure and the accumulative 
closure, and also provide audiovisual 
warnings when the closure rate reached a 
preset critical value. 



109 



ACKNOWLEDGMENTS 



The author wishes to thank Bob 
Ochsner and Tom Heaps of SUFCO for 



their help in testing the closure-rate 
instruments. 



MECHANICAL INSTRUMENT 



DESCRIPTION 

A mechanical closure-rate instrument 
was built first. It consists of two 
rugged telescoping potentiometric exten- 
someters and a digital readout-control 
box (fig. 1). The extensometer (fig. 2) 
is designed to accommodate a height of 
from 4-1/2 to 12 ft with a measurement 
It is spring-loaded over 
Long (100- to 125-ft) 
cables connect the extensometers to the 
readout box, permitting the operator to 
remain in a safe, supported area. A 
breakaway feature on each extensometer 
allows it to be pulled from the fall area 
by its electrical cable. 



range of 6 in. 
this 6-in range. 



The operator watches the digital read- 
out on the control box (fig. 3) during 
mining of the pillar. When a predeter- 
mined critical rate of closure is indi- 
cated, he retrieves the extensometer 
(fig. 4) and signals the miner operator 
to pull back. The control box (fig. 5) 
provides two digital visual readouts. 
One display shows the closure rate, the 
other total accumulative closure from 
time zero. The operator can preset any 
closure rate from to 1 in/min. When 
the preset rate is reached, an alarm 
light illuminates and an audible alarm is 
sounded. The control box can monitor two 
extensometers individually or alter- 
nately. The system is battery-operated. 




FIGURE 1. - Jvlechanical closure-rate system. 




FIGURE 2. - Extensometer. 



110 




FIGURE 3. - Control box, on-site. 



conqjletely portable, and can be quickly 
moved from one place to another. 

The closure-rate instrument, though 
primarily designed for retreat mining, 
can be used for any activity where knowl- 
edge of roof-to-floor closure rate or 
total displacement is required. The sys- 
tem provides a 0- to 6-in measurement 
range with 0.1 pet accuracy for both rate 



and total closure. Any extensometer in- 
stallation displacement can be zeroed out 
with zeroing potentiometers. This zero 
value can be recorded and reset if 
multiple extensometers (greater than two) 
are used. The auto alarm can be set to 
any closure-rate value from to 1 in/ 
min, with a 0.01-in resolution and 0.1 
pet accuracy. 



HI 




FIGURE 4. - Installed extensometer. 












CIRCUIT DESCRIPTION 

Figure 6 is a timing diagram for the 
operation of the electronic measurement 
circuits. Because of the small changes 
in actual rate, normally up to about 0.2 
in/min, it was decided to use a 10-s sam- 
ple period for greater accuracy. The ex- 
tensometer is sampled at the beginning 
(S2) and at the end (S,) of a 9-s period 
by sample and hold circuits. The two 10- 
ms samples are compared, and any differ- 
ence is converted to a directly propor- 
tional frequency, counted for 100 ms , and 
displayed as rate in inches per minute. 
A total read cycle consists of taking the 
two samples, resetting the counter, read- 
ing the value, and displaying it dig- 
itally. The extensometers are auto- 
matically measured alternately or can be 
continuously monitored on a singular bas- 
is. The extensometer circuit (if in 
auto) will automatically cycle to extens- 
ometer No. 2 with the extensometer toggle 
switch in auto position. When No. 2 has 
been sampled and displayed. No. 1 extens- 
ometer is automatically toggled back into 
the circuit. 



FIGURE 5. - Control box, closeup. 



Figure 7 is an electrical diagram 
of the readout-control box. Each 



112 



1 Hz dispiay 



K 



Time zero 



1 



J 



1 Hz clock .JoUUJIUIUTLfBUeUTLJsLJ^^ 



Sj sample 
S, sample 
Reset counter 
Read 



il 



10 ms 



il 



10 ms 



il 



10 ms 



Jl 



10 ms 



Toggle ext H ^^ "^^ 

F/F 



U 100 ms 
[-[10 ms 



Note: Timing as follows 

1. Extensometer 1 is selected by "Toggle ext pulse." 

2. Extensometer voltage is sampled "Sz" and is held to be compared 
with Si. 

3. Si is sampled and compared with Sz- 

4. Counter is reset ready to count. 

5. Counter reads V to F count. 

6. Count displayed for 5 s. 

7. Extensometer 2 is selected by "Toggle ext pulse." 

8. Same sequence of reading for Extensometer 2. 

9. Extensometer 1 is selected by "Toggle ext pulse." 

FIGURE 6. - Timing diagram. 



extensometer electrical output is fed Accuracy: 
directly to a high-input impedance ampli- 
fier. The outputs from these amplifiers Resolution: 
are routed to a field-effect transistor 
switch, which selects the extensometer to Alarm: 
be measured. One route converts the ex- 
tensometer voltage to frequency and is 
displayed as inches of total accumulative Readout: 
displacement. The same voltage is routed 
to the sample-and-hold circuitry, where 
it is sampled, summed, and converted to 
rate in inches per minute. This differ- 
ence, or summed voltage, is also sent to Power: 
the comparator for the alarm function. 

INSTRUMENT SPECIFICATIONS 

Measured range: to 6 in displacement 

Closure rate: to 6 in/min 



+0.1 pet 

0.001 in 

to 1 in/min with 
0.01-in resolution 

Digital LED display: 
two; one for total 
displacement; one 
for closure rate. 

Battery power 12-h 
capacity. Intrinsi- 
cally safe (MSHA 
approved) . 



113 



Ext. No. 1 



»- 6.9 V regulated 




V/F, 



SW1 



Ext. No. 2 



>- 6.9 V regulated 




FET 
SW 



Toggle 




Voltage 
to f req 



Counter 



' Display \ 



jf Warning 
< — 1^ ''9'^* ^ !► Alarm 



Total closure 



^ 



Audible 
alarm 



J-*fset 



Cpl 



INVn 



osc 



Auto- 
Dlvider circuits 



Sample 

and 
hold S, 



VF, 



Voltage 
to freq 



100 
Hz 



10 



10 



Hz 



-MO 



Hz 




Closure rate 



Timing circuit 

100 ms 
— ► Read 

►lO ms 

Sample 

FIGURE 7. - Block diagram of closure-rote instrument. 
ULTRASONIC INSTRUMENT 



A nonobstructing method of measuring 
roof-to-floor closure has always been 
desirable in underground ground-control 
measurements. The mechanical extensom- 
eter has several drawbacks, such as high 
cost and being hard to use in high- 
traffic areas. Thus the Bureau undertook 
a project to determine the feasibility of 
using ultrasonics for convergence mea- 
surement underground. The new Polaroid 
sonar camera transducer and ranging sys- 
tem became a prime candidate. A small 
ultrasonic transducer of this type could 
provide an inexpensive, nonobstructing 
means of measurements. 

The Bureau has thus far developed a 
small, yet inexpensive, handheld ultra- 
sonic unit, shown in figure 8, that can 
measure distances from 1 to 35 ft ±0.02 
ft. The unit is excellent for general 
survey work, measuring high roofs, etc. 



A small, portable closure-rate measuring 
device was also built with an ultrasonic 
remote transducer that can be placed up 
to 100 ft or more from the readout in- 
stmment. This unit is shown in figure 
9. This instrument reads both distance 
as well as rate of closure. It has a 
measurement range of 1 to 35 ft ±0.01 ft 
with 0.001-ft resolution. The first 
reading displayed is the distance. The 
rate is displayed 6 s later. A thumb- 
wheel switch allows for setting an alarm 
limit for rate of closure. A visual and 
audible alarm is provided. The alarm 
limit can be set from to 9.9 ft/min. 

Because the ultrasonic instrument is 
still in the design and testing phase, 
and since permissibility approval has not 
been received from MSHA, its electrical 
schematics and operational information 
will be made available at a later time. 



114 




FIGURE 8. • Portable handheld ultrasonic unit. 



115 




FIGURE 9. - Portable ultrasonic closure-rate instrumento 



FIELD TESTS 



In June 1981, the mechanical closure- 
rate instrument was placed for field 
testing in SUFCO's No. 1 Mine (figs. 3-4) 
near Salina, UT. The test results have 
exceeded expectations, and roof-caving 
predictions have proven very accurate. 
In most cases, the instrument has warned 
of an impending roof -caving within min- 
utes of the event. It was reported that 
use of the closure-rate instrument has, 
in general, led to increased coal recov- 
ery and productivity. 



The readout instrument, except for a 
fuse blown during battery replacement, 
has been trouble-free. A design change 
in the breakaway mechanism and substitu- 
tion of heavier gauge extensometer rods 
are the only modifications made to the 
system thus far. It is noteworthy that, 
to date, no extensometer s have been lost 
during caving. 



116 



CONCLUSIONS 



With 22 months of field tests we con- 
clude that the mechanical closure-rate 
Instrument appears to be a viable tool 
for roof hazard prediction In retreat 
mining operations once the critical clo- 
sure rate has been determined for a par- 
ticular mining location. Most retreat 



cycles. In general, have Increased ton- 
nage per production shift. 

Preliminary test results of the ultra- 
sonic closure-rate unit appear excellent. 
It Is hoped that this unit will also Im- 
prove safety underground. 



117 



GROUND INSTALLATION EQUIPMENT 

REMOTE MANUAL ROOF BOLTERS 
By John E. Bevani 



ABSTRACT 



Coal mine accident statistics show that 
18 pet of all roof-fall fatalities in- 
volve roof bolting. This is 30 pet 
higher than for any other occupation 
category. Industry analyses show that 
roof bolters were involved in 15 pet of 



lost-time accidents. Rapid placement of 
permanent roof support appears essential 
to safety as well as long-term roof sta- 
bility. This paper investigates one 
method for placement of permanent roof 
support. 



BACKGROUND 



The MESA report, "Analysis of Fatal 
Roof-Fall Accidents in Coal Mines, 1972- 
1975," indicates that 18 pet of all roof- 
fall fatalities involved roof-bolter 
operators and helpers. This group of 
workers experienced 30 pet more fatal- 
ities than any other occupation category. 

The report "Injuries Associated With 
Roof or Rib Bolting and Bolting Machines 
in Underground Coal Mines, 1978-1982" 
analyzed 5,777 underground coalmine bolt- 
ing or related bolting machine injuries 
reported to HSAC from 1979 through 1982. 
The results were — 

1. Drilling of the roof or rib 
accounted for 2,457 injuries (45 pet). 

2. Installation of bolts in roof or 
rib~l,634 (28.3 pet). 

3. Tramming the machine — 727 (12.5 
pet). 



4. Unknown (due to insufficient data 
to classify) — 959 (16.6 pet). 

Industry analyses of lost-time acci- 
dents are not complete, but the following 
figures are typical. One company's 
Safety Department reports that roof-bolt 
operators were involved in 15 pet of 
their lost-time accidents. Of the roof- 
bolting accidents, 19 pet were caused by 
roof falls , and 48 pet were caused by 
being struck or caught by rotating tools. 

The above information indicates the 
danger involved with roof bolting and 
roof control. Many other non-roof-bolter 
injuries and fatalities are also a result 
of insufficient roof support. Rapid 
placement of permanent roof support also 
appears to be beneficial to mid- and 
long-term roof stability as well as imme- 
diate safety. 



RATIONALE 



The Bureau of Mines has addressed 
these problems through two areas of re- 
search: (1) by removing the bolter oper- 
ator from the immediate dangers of bolt 
installation; and (2) by providing means 
for a safer and more timely bolting 
system. 

'Mechanical engineer, Spokane Research 
Center, Bureau of Mines, Spokane, WA. 



A roof-bolt inserter (RBI) developed 
allowed a longer-than-seam-height bolt to 
be installed by bending the horizontally 
carried bolt into the vertical orienta- 
tion of the mine roof-bolt hole. A 
longer-than-seam height (LTSH) drill or 
flexible roof drill developed by con- 
tract allowed long holes to be drilled 
in low coal. The RBI and LTSH drill add 
flexibility of package design, which 



118 



allows the operator to be removed from 
the immediate bolting area, under sup- 
ported roof and away from the dangers of 
rotating and moving equipment. The oper- 
ator can be placed outby the last perman- 
ent row of roof support and/or under can- 
opy protection. Here the operator can be 
protected while still using his or her 
abilities. 

An automated miner-bolter to use the 
concept of mounting the two systems on a 
continuous miner would allow one-pass or 
truly continuous mining. 

However, many problems of an automated 
system may be caused by the substitu- 
tion of the operator's function with 
mechanical-electrical-hydraulic systems 
and artificial intelligence. Manual dex- 
terity, memory, logic, audio, visual, and 
sensory capabilities of the operator have 
been replaced in an attempt to remove him 
or her from the dangers of the immediate 
bolting station. It appears that any 
viable system must be greatly simplified; 
hence, replacing some of the capabilities 
of the operator is unjustified and unwar- 
ranted if the operator can supply these 
functions while being protected. 

Six concepts featuring hands-off drill- 
ing, remote control and/or automated 



sequencing, and improved productivity 
features were developed. From these six 
concepts and initial designs, the artic- 
ulated remote manual roof -bolter (ARM 
bolter) and the remote-manual bolter 
(REM bolter) were chosen for further 
development. 

The ARM bolter (fig. 1) is a roof 
bolter for low seams , which enables an 
operator to perform bolting functions 
while under a canopy protection and per- 
manent roof support. The operator sits 
in a reclined position in a cab with his 
head approximately 8 ft from the bolt- 
hole location. From this location, the 
drilling, bolting, and torquing of bolts 
are controlled. Bolts ranging in length 
from 4 to 8 ft are fed into the bolter 
component assembly. The machine's over- 
all tram height is approximately 33 in 
and it is capable of installing bolts in 
seams ranging from 37 to 60 in. The lim- 
itation in seam height is due to bolter 
design and not necessarily limited by 
conceptual considerations . 

The heart of the ARM bolter is the 
bolter component assembly. This assembly 
houses a flexible roof drill, roof -bolt 
inserter (RBI), torque thrust assembly, 
plate magazine, feed and receive mechan- 
isms, and a component carriage to house 



I 




FIGURE lo - Articulated remote manual (ARM) roof bolter. 



119 



the above elements. The bolter component 
assembly interfaces with the front end of 
the articulated vehicle and is raised and 
lowered by means of two elevation assem- 
blies. The operator manually assembles 
the mechanical bolt and anchor, then 
loads the assembled bolt (less anchor 
plate) into the feed and receive mechan- 
ism prior to drilling the hole. The 
anchor plates are loaded into the plate 
magazine at the same time. The operator 
enters the operator station (under canopy 
support) and trams the bolter to the 
proper bolting position as dictated by 
the mine plan. The bolter component as- 
sembly is raised to the roof, and the 
bolt hole is drilled by the longer- 
than-seam-height drill. After the hole 
is drilled to the desired length, the 
drill string is retracted and the drill 
is indexed away allowing the RBI-torquer 
assembly to be indexed to align with the 
drilled hole. The plate feed assembly 
installs the plate on the roof bolt, then 
the RBI installs the bolt into the hole. 



The RBI expands allowing the torquer to 
engage, insert, and torque the bolt as- 
sembly. The RBI and drill are indexed to 
their stow position, the bolter component 
assembly and roof jacks are lowered, and 
the ARM bolter is ready to tram to the 
next bolt installation. The control is a 
combination of air logic control and re- 
mote operator control. Remote operator 
control is available in the event of an 
air logic system failure. These opera- 
tions, done during bolt installation, are 
achieved while the operator is under can- 
opy and permanently supported roof. 

The second concept being developed 
(fig. 2) by the Bureau is the remote man- 
ual bolter (REM bolter). This concept 
uses items from the longer-than-seam- 
height drill program, but previous prob- 
lems with automated modules directed 
attention toward enchancing, rather than 
replacing the operator. Many of the same 
design criteria used for the ARM bolter 
were used in the development of the REM 




FIGURE 2. . Remote manual (REM) roof bolter. 



120 



bolter. The operator is used more in the 
REM bolter concept than the ARM bolter 
(attempting to make a simpler system). 
The same steps are required to install 
the bolt, but here the operator is re- 
sponsible for assembling the complete 
bolt. Any anomaly must be compensated 
for by the operator. 

The REM bolter is trammed into posi- 
tion, and the drilling is initiated. The 
RBI is mounted on a track which allows it 
to slide back near the operator where the 
assembled bolt is manually placed into 
the RBI. When the drilling cycle is com- 
pleted, the drill is indexed to its 
stowed position, and the RBI is run for- 
ward on the track and indexed under the 
drilled hole. Final positioning of the 
RBI, if required, can be done by moving 
the bolter or moving the arm that carries 
the RBI. The RBI then inserts the bolt 
into the hole and is indexed to its stow 
position. The torquer is indexed into 
position and coupled with the bolt. The 



bolt is torqued, the torquer is stowed, 
the floor jack is released, and the 
bolter is ready to be trammed to the next 
bolting position. The operator has not 
moved from his protected position. 

All control functions of the REM bolter 
are operator-controlled. No sensory 
feedback or logic circuit, lockout, etc., 
is employed. This is done for simplicity 
and its associated reduction in cost, 
weight, and size. 

The future for remote bolters should 
hold great promise. The remote-manual 
concept can easily be adapted to resin 
bolts or a combination of mechanical 
and resin. Water-jet-assisted flexible 
drills could increase the number of roof 
conditions where the bolter can be used. 
(Water-jet-assisted flexible drills may 
replace rotary-impact drilling now used 
on severe roof conditions). New concepts 
like the inorganic grout injection de- 
vices could be easily adapted. 



RESULTS 



A 4-month underground test of the REM 
bolter began February 25, 1983, at a mine 
near Daisytown, PA. A soft band of shale 
in the mine roof resulted in flex drill 
problems. The dust system appeared to be 
plugging, which resulted in impacting the 
drill string in the hole. Another prob- 
lem area was the operator's difficulty in 
inserting the bolt and coupling the 
torquer to the bolt head after it was 
inserted. 

Another working section was made avail- 
able by mine personnel and initial drill 



problems did not reoccur. Bolt installa- 
tion of 50 holes per shift has been 
achieved. 

The ARM bolter has been tested for 
a total of 5 weeks in a West Virginia 
mine. A total of 163 bolts have been 
installed. No major problems were en- 
countered; however, some correctable 
problems have been encountered with the 
drill, cycle time, and dust collection 
system. 



CONCLUSIONS 



Both the ARM bolter and the REM bolter 
address the same problems, but the path 
taken by each is different. Some of the 
questions to be answered are: 



3. Can the operator find the hole to 
insert the bolt, and can it be done in 
varying seam heights with undulating top 
and /or bottom? 



1. What degree of automation or semi- 
automation is required? 

2. Can the operator develop the 
needed skills to couple and uncouple the 
torquer to the bolt or can a control sys- 
tem do it better? 



4. Will the operator be fatigued or 
will active participation result in a 
safer, more alert operator? 

The program goal for the REM and ARM 
bolters was to develop a concept that 



121 



will protect the operator , be econom- 
ically feasible, and be implemented by 
industry into a production machine. The 
program development and subsequent test- 
ing has shown the concept as viable. 
Automated bolters were costly and com- 
plex. The remote bolters are less expen- 
sive and much simpler while maintaining 
operator safety and high production 
levels. 



Future generation machines will be more 
efficient and probably less costly. The 
degree of automation that can be effec- 
tively used in the commercial underground 
mining environment will increase as the 
state-of-the-art in robotics, controls, 
software, etc., increases, but any suc- 
cessful machine must be designed around 
the human operator, still the most impor- 
tant element of any system. 



122 



FIELD TEST OF AN AUTOMATED TEMPORARY ROOF SUPPORT (ATRS) USED ON A LOW-COAL, 
SINGLE, FIXED-HEAD ROOF BOLTING MACHINE (SQUIRMER) 

By Charles T. Chislaghi^ and Thomas E. Marshall^ 



ABSTRACT 



An economical, remotely operated (auto- 
mated), temporary roof support (ATRS) has 
been developed by the Bureau of Mines for 
use on a single, fixed-head roof bolting 
machine (squirmer) that operates in low- 
coal seams (<42 in thick). This ATRS 
eliminates the need for workers to go 
under unsupported roof to set or remove 
temporary support prior to or during the 
roof bolting cycle — a task that annually 
accounts for approximately 20 pet of all 
roof fall fatalities. It can be adapted 



on any squirmer used in the U.S. low-coal 
fields. A prototype ATRS, designed and 
built at the Bureau's Pittsburgh Research 
Center, was field-tested at Imperial Col- 
liery Co.'s Mine No. 20 in Eskdale , WV. 
The Mine No. 20 amended roof control 
plan, which requires the use of the Bu- 
reau's ATRS as temporary support during 
face bolting, has been approved by the 
Mine Safety and Health Administration 
(MSHA) . 



INTRODUCTION 



A statutory provision of the Federal 
Coal Mine Health and Safety Act of 1969 
states that "No person shall proceed be- 
yond the last permanent support unless 
adequate temporary support is provided."-^ 
However , since the time the law was 
written, there have been no means avail- 
able for squirmer operators and helpers 
to set temporary supports from under per- 
manently supported roof. This provision 
was interpreted to mean "In areas where 
permanent artifical support is required, 
temporary support should be used until 
such permanent support is installed,"'* 
and "Only those persons engaged in in- 
stalling temporary support should be 
allowed to proceed beyond the last per- 
manent support until such temporary sup- 
ports are installed."^ Annually, ap- 
proximately 20 pet of all roof-fall 
fatalities involve miners who have gone 
beyond the last permanent support to set 
or remove temporary roof support prior to 
or during the roof bolting cycle. 



'Mining engineer. 
^Engineering technician. 
Pittsburgh Research Center, 
Mines, Pittsburgh, PA. 
^30 CFR 75.200. 
"^SO CFR 75.200-13(a){1 ) . 
^30 CFR 75.200-13(a)(2) 



Bureau of 



Because of space limitations in low 
coal, not many ATRS's have been commer- 
cially developed for squirmers, although 
many different ATRS systems have been 
commercially developed for roof bolting 
machines used in high coal. Most ATRS's 
designed for squirmers create a situation 
that reduces or compromises the existing 
safety level, with a greater safety haz- 
ard to squirmer operators and helpers 
working inby the last row of permanent 
support. 

Over 3,500 squirmers are in use today 
in southern West Virginia, eastern Ken- 
tucky, and southwestern Virginia, and 
approximately 60 pet of these have no 
ATRS, cab, or canopy. Moreover, low-coal 
mine operators and owners in West Vir- 
ginia have a need for ATRS because West 
Virginia mine law requires that roof 
bolting machines used in working places 
of West Virginia coal mines be equipped 
with ATRS, regardless of coal seam 
height. 

All design work and prototype fabrica- 
tion was done by the Roof Support Group 
at the Pittsburgh Research Center. All 
fieldwork was done in the No. 2 Gas seam 
(36 to 42 in thick) at Mine No. 20 of the 
Imperial Colliery Co. in Eskdale, WV. 



123 



ACKNOWLEDGMENTS 



The Bureau acknowledges the coopera- 
tion it received from personnel of the 
Imperial Colliery Co., MSHA Mt. Hope 
Subdistrict, and MSHA Bruceton Safety 



Technology Center. Without their techni- 
cal suggestions and assistance, this 
project could not have been completed. 



DESCRIPTION OF ATRS 



The Bureau of Mines ATRS is based on a 
modified and improved Lee Engineering 
design for a squirmer ATRS. It consists 
of a 10-ft-long, steel, wide-flange beam 
supported by two double-acting, telescop- 
ing hydraulic cylinders (fig. 1). A 
steel sleeve, mounted on the bottom cen- 
ter of the beam, is designed to fit over 
the top of the squirmer drill head. The 
ATRS is carried from place to place and 
row to row on the squirmer drill head, 
but is not an integral part of the 
squirmer during bolting. During bolting 
it is connected to the squirmer only by 
two hydraulic lines. Because the ATRS 
only weighs about 400 lb, the squirmer 



drill head and boom do not have to be re- 
built to carry it. Total cost of the 
beam and cylinders is approximately 
$1,800. In-house fabrication of the ATRS 
took 8 worker-hours. 

The Bureau's ATRS design meets MSHA's 
general design requirements and West Vir- 
ginia's design and operating requirements 
for such support. Both hydraulic cylin- 
ders supporting the ATRS have check 
valves to prevent sudden collapse of the 
ATRS in the event of a ruptured hydraulic 
line or broken hydraulic fitting. In 
addition, the ATRS hydraulic circuit con- 
tains an accumulator , charged by squirmer 




FIGURE 1. - Bureau of Mines ATRS. 



124 



line hydraulic pressure, which keeps the 
ATRS firmly set against the mine roof 
even if the roof rock is pulled up dur- 
ing the bolting cycle. The ATRS can 



elastically support the minimum required 
deadweight load of 33,750 lb; this 
capacity is certified by a professional 
engineer. 



SQUIRMER STREAMLINING 



Imperial Colliery personnel streamlined 
a 15-yr-old FMC model 300 squirmer for 
the field test. West Virginia State Mine 
Law requires the streamlining of any roof 
bolting machine before it can be retro- 
fitted with ATRS. The ATRS controls were 
located 5 ft back from the drill head so 
that they can only be operated from be- 
neath permanently supported roof. Inch 
tram controls were located at the drill 
station, and inch tram speed was reduced 
to 65 ft/min. Full tram controls were 
located with the ATRS controls, and the 



full tram speed was left at 150 ft/min. 
No ATRS controls were located at the 
drill station. Other streamlining work 
included removal of the bolt tray and 
tram deck, installation of low-value, 
high-torque tram motors, and moving the 
squirmer front wheels 8 in forward to 
provide space for the ATRS and full tram 
controls. Total cost of this work, which 
required 96 worker-hours, was $5,500. 
The Bureau piped the drill circuit to the 
ATRS and piped the ATRS circuit on beam 
at a cost of $150 and 32 worker-hours. 



FIELD TEST AND RESULTS 



With MSHA and West Virginia approval. 
Imperial Colliery placed the ATRS in the 
production cycle at Mine No. 20 for 5 
months. Bolting was on 4-ft center, with 
20-ft-wide entries and crosscuts. The 
cycle was the following: 

Step Description 

1 The squirmer operator, at the full 
tram controls, trams into the cen- 
ter of a place and stops when the 
ATRS is under the last row of per- 
manent support. 

2 The operator lowers the drill head 
(and ATRS) to the mine floor using 
a boom control located beside 
the full tram and ATRS controls; 
moves from the full tram position 
to the beam (ATRS); and unhooks 
the hydraulic cylinder leg on the 
operator side that is chained to 
the beam, while the helper does 
the same to the leg on the right 
side. 

3 The operator raises the drill head 
(and ATRS) , using a boom control at 
the drill station, just high enough 
to let the legs hang down without 
scraping the mine floor; locks the 
legs perpendicular to the mine 
floor; moves back to the full tram 
and ATRS controls; and trams the 



squirmer inby without the legs 
scraping mine floor or the beam 
scraping mine roof. 

The operator stops when under the 
last row of permanent support. The 
ATRS is now 5 ft inby the last row 
of permanent support and 5 ft from 
each rib. 

The operator places the ATRS 
against the roof using the boom 
control located beside the full 
tram and ATRS controls, and then 
lowers the legs to the mine floor 
using the ATRS control until the 
beam is firmly set against the roof 
and the legs are firmly set against 
the mine floor. 

The operator lowers the drill head 
away from the beam using the boom 
control located beside the full 
tram and ATRS controls. 

At this point, the operator moves 
to the drill station, pushes in the 
diversion valve which diverts all 
hydraulic fluid from the full tram 
circuit to the inch tram circuit, 
and "inches" the squirmer to the 
left rib to begin bolting. During 
bolting the squirmer is connected 
to the ATRS by only two hydraulic 
lines. 



125 



8 After a row of permanent support is 
installed, the operator raises the 
drill head into the beam using the 
boom control at the drill station; 
pulls out the diversion valve which 
diverts all the hydraulic fluid 
back to the full tram circuit from 
the inch tram circuit; moves to the 
full tram and ATRS controls; and 
raises the legs using the ATRS 
control. 

9 The operator lowers the drill head 
(and ATRS), using the boom control 
located beside the full tram and 
ATRS controls , just enough to tram 
the squirmer inby without the legs 
scraping mine floor or the beam 
scraping mine roof. 

10 When under the row of permanent 
roof support that has just been in- 
stalled, the operator stops and re- 
peats steps 5 through 9. This 
cycle is repeated until the last 
bolt is in place. 

11 Then the operator raises the drill 
head into the beam using the boom 
control at the drill station; un- 
locks the legs ; pulls out the di- 
version valve; moves to the full 



tram and ATRS controls; raises the 
legs using the ATRS control; moves 
back to the drill station; lowers 
the drill head (and ATRS) to the 
mine floor using the boom control 
at the drill station; chains the 
leg, on the operator side, to the 
beam while the helper does the same 
to the leg on the right side; moves 
back to the full tram and ATRS con- 
trols; turns the squirmer 180°; and 
trams to the next place where steps 
1 to 11 are repeated. 

No operating or maintenance problems 
were encountered during the 5 months of 
testing. With the addition of the ATRS, 
the squirmer could still turn 180° within 
the 20-ft-wide entries and crosscuts and 
could tram through check curtains and 
line brattice without pulling them down. 
Comparative time studies of the same 
bolting crews showed that it took an av- 
erage of 5 min less to bolt a place with 
the ATRS than with mechanical jacks. The 
bolting crews preferred the ATRS. After 
testing, an amended roof control plan re- 
quiring the use of the Bureau's ATRS dur- 
ing face bolting at Mine No. 20 was sub- 
mitted by the Imperial Colliery Co. and 
approved by MSHA — District 4, Mount Hope 
Subdistrict, Montgomery Field Office. 



CONCLUSIONS 



The Bureau's ATRS eliminates the need 
for squirmer operators and helpers to go 
under unsupported roof to set or remove 
temporary support prior to or during the 
roof bolting cycle. The squirmer oper- 
ator will always be under permanently 
supported roof while setting or removing 
the ATRS and will not be able to bolt 
inby the ATRS because of its control 
location. The Inexpensive and light- 
weight ATRS does not reduce the squirmer 
operator's work space. It has the poten- 
tial to be immediately used in some 70 
pet of U.S. low-coal mines. Although the 
Bureau's ATRS was field tested only with 



the FMC model 300 squirmer , it can be 
adapted to the drill head of any squirmer 
operating in low coal and it can be 
fabricated in any mine shop. If a 
streamlined squirmer is available, the 
ATRS can be retrofitted to the squirmer 
during maintenance shifts , and if prob- 
lems occur with the ATRS during the bolt- 
ing cycle, it can be disconnected from 
the squirmer to allow bolting to continue 
with mechanical jacks. It has the poten- 
tial to eliminate roof fall fatalities 
and injuries and may lead to increased 
productivity. 



126 



ROOF SUPPORT SYSTEMS 

DEVELOPMENT OF EPOXY GROUTS AND PUMPABLE BOLTS 
By Robert R. Thompson'' 



ABSTRACT 



Good roof control is critical to the 
coal mining industry. The Bureau of 
Mines recognized the need for a remotely 
placed roof bolt of noncorrosive materi- 
als, which could be remotely installed 
in longer -than-seam-height lengths. The 
system developed has a fiberglass core. 
The core is made up of four 1/4-round 
sections, which can be coiled on the 
machine for storage and then formed into 
a hollow core during insertion into the 
drilled hole. The adhesive used is a 



fast-setting epoxy resin. The resin was 
tested underground in cartridge form. 
The test results showed bonding strengths 
greater than those available with the 
polyester resin now being used. Core 
handling and pumping equipment was de- 
signed, built, and placed on an existing 
roof bolter. The equipment is now being 
laboratory-tested and will be tested 
underground in a coal mine using a stan- 
dard MSHA-def ined, two-intersection test. 



INTRODUCTION 



Personnel safety is the first benefit 
of good roof control, but productivity 
can also be affected significantly by im- 
proved roof control procedures. Most of 
the easily mined coal has been extracted, 
and future mining will take place in 



areas with difficult roof-control prob- 
lems . The Bureau of Mines recognized the 
need for a remotely placed roof bolt 
of noncorrosive materials that could be 
remotely installed in longer-than-seam- 
height lengths . 



BACKGROUND 



The system envisioned would consist of 
a hollow fiberglass-reinforced bolt and 
an adhesive that would be pumped into the 



annulus between the bolt and rock. The 
Bureau initiated work to develop such a 
system in 1978. 



EARLY DEVELOPMENT 



During the early part of the work, pol- 
yesters, urethanes, acrylics, epoxies, 
and inorganics (cement and gypsum) were 
examined for use as an adhesive. Poly- 
ester resins widely used today in roof 
bolting are supplied in cartridges that 
are used with steel bolts. The poly- 
esters did not lend themselves to pumping 
because of resin instability and low 
flashpoints. Urethanes and acrylics were 
easily pumped and had fast set times, but 
were either too expensive or could not 

'Research structural engineer, Spokane 
Research Center, Bureau of Mines, Spo- 
kane , WA . 



meet adhesive strength requirements. A 
coal-tar-based epoxy, one part resin to 
one part hardener , was developed which 
could be mixed in a static mixer and 
pumped by conventional liquid-handling 
equipment. Gel time requirements of the 
epoxy were met by preheating of the com- 
ponents. Cements had fast set times but 
unacceptable mixing and pumping charac- 
teristics. It was determined that the 
epoxy and cements were the most promising 
and would require additional laboratory 
testing during the second phase of the 
program. 



127 



The core shape 
that the core 



Also, during the initial phase of the 
program, several shapes of fiberglass 
bolt cores were examined, 
requirements were such 
could be coiled on reels for storage on 
the machine, then formed into a hollow 
core bolt during installation. Surface 
modifications on the smooth fiberglass- 
reinforced polyester (FEIP) core to in- 
crease bonding strengths of the adhesives 
appeared necessary. A variety of special 
wraps and cloth meshes, fabricated on the 
outer surface of the smooth cores, delam- 
inated under the severe pull strengths 
used to evaluate the system. The best 
method found was to cut a diagonal groove 
into the outer surface. This provided 
the mechanical interlock necessary to 
achieve the required strengths. 



Initial designs led to two FRP bolt 
core shapes for laboratory testing. 
An axially pleated, flat-sheet material 
was tried. It could be stored in a 
flat coil and pulled as a "tape" to 
the placement head, which would fold 
it to form a hollow, cylindrical core. 
The second core configuration was 
based on discreet lengths of FRP, each 
a 1/4-round section which could easily 
be bent at the head, and the four pieces 
formed into a hollow core prior to in- 
sertion into the drilled hole. Again, it 
was determined that both bolt core con- 
cepts required additional laboratory 
testing during the second phase of the 
program. 



PHASE II DEVELOPMENTS 



The second phase of the program per- 
mitted more extensive laboratory testing 
of the core and grout candidates. Con- 
crete blocks and test equipment were con- 
structed to allow grouting and pull test- 
ing of both FRP bolt cores with epoxy and 
cement adhesives. 

With the aid of several chemical com- 
panies, providing their own funding, 
inexpensive hardeners were developed 
which produced fast gel times without 
need for preheating. Additional labora- 
tory work with the cements still produced 
unacceptable mixing and pumping. It was 
determined that the epoxy showed the most 
promise and that work with the cements 
would be terminated. 

A series of four epoxy formulations was 
developed that met the 1-min gel times 
and early strengths needed over the re- 
quired temperature range of 40° to 



100° F. They had a 1:1 ratio of resin to 
hardener, were highly filled, and easily 
mixed within a static mixer. A twin cyl- 
inder, positive-displacement piston pump 
was used to meter, mix, and dispense the 
epoxy resin. 

Both the flat-sheet and the segment 
core designs were adaptable for continu- 
ous insertion of longer-than-seam-height 
lengths. The folded sheet core tended to 
collapse on itself under cross-shear, 
while the segmented core design remained 
stable and was stronger in cross-shear. 
It was therefore decided to use the seg- 
mented core design. 

Laboratory tests are continuing to ver- 
ify that the segmented core design and 
the new epoxy systems meet or exceed all 
the strength requirements recommended by 
the Bureau. 



EPOXY CARTRIDGE DEVELOPMENT AND FIELD TEST 



Evaluation of epoxy grouts in mine con- 
ditions bypassed the pumping system de- 
velopment by packaging in cartridge form. 
Steel bolts with epoxy cartridges were 
installed in several mines under condi- 
tions similar to those used for polyester 
cartridges. In this way, fully grouted 



roof bolts using the epoxy formulation 
could be easily compared to existing sup- 
port systems. 

Pull tests on roof bolts with 1- 
ft point anchors were conducted off- 
section in the mine. Underground results 



128 



verified laboratory testing, in that the 
epoxy bolts could be installed as easily 
as the polyester and with greater bonding 
strengths. 

A double intersection roof-bolt test 
was carried out in a freshly mined area 
(pretimbered) of a working Eastern coal 
mine. Over 200 epoxy-resin-grouted steel 
bolts were installed in the test without 
any installation problems. After a 



suitable return passage had been mined, 
the timbers were removed and the roof was 
evaluated for stability. The results 
indicated that the bolted strata were 
successfully supported and that the epoxy 
cartridges could be used as a viable 
alternative to the polyester system. 
Several private con^janies are pursu- 
ing the possible marketing of epoxy 
cartridges. 



NEW PROGRAM 



As a result of the laboratory and mine 
evaluations of the epoxy bolting system, 
the Bureau entered into a cost-sharing 
research program to develop the necessary 
mineworthy installation equipment to 
evaluate the concept of a remote- 
ly placed, pumpable, longer-than-seam- 
height, epoxy FRP, noncorrosive, roof- 
bolt system. 

The contractor is furnishing a bolter 
chassis for the test period. Four major 
subsystems must be added to the chassis 
for placement of the pumpable roof bolts: 
the pumping equipment, purge system, 
fiberglass core-forming equipment, and a 
placement head for interfacing each of 
the subsystems with the drilled roof-bolt 
holes. 

The contractor will incorporate, on the 
bolter, all the new design features shown 
to be needed. The compact pumping system 
design enables easy material loading and 
repair from the rear of the bolter. The 
resin will be pumped by two piston pumps 



which are driven by a cylinder mounted 
between the pumps. The system also in- 
cludes a static mixer, valves, and 
hoses. 

Laboratory testing indicates that a 
high-pressure inexpensive water purge 
completely cleans the static mixers. 

The bolter system design includes a 
three-segment 3/4-in-diam core, three- 
reel storage, and handling system. The 
three segments of core will be pulled 
from the storage reels and driven through 
rollers designed to form a 3/4-in hollow 
core bolt when the three segments arrive 
at the drilled hole. 

The placement device is designed to 
bring together the epoxy dispenser, purg- 
ing system, and core former into a conmion 
head where the components can be prepared 
for roof bolting. This head will inter- 
face directly with the mine roof to pro- 
vide multiple composite bolt placement. 



FUTURE PLANS 



During the next 4 months , the equipment 
will be laboratory tested. The results 
will be reviewed and systems redesigned 
as needed to ensure a meaningful field 
evaluation. 



involved are satisfied with the system, a 
standard MSHA-def ined, two-intersection 
test will be conducted in a cooperating 
coal mine. 



Field evaluation will be conducted off- 
section in a coal mine. When all 



129 



DEVELOPMENT OF LIGHTWEIGHT HYDRAULIC SUPPORTS 
By John P. Dunfordi 



ABSTRACT 



The Installing of temporary roof sup- 
port is an integral part of underground 
coal mining. The most common forms of 
temporary supports are wooden posts and 
metal jacks. Both wooden posts and non- 
yielding steel jacks are heavy and cum- 
bersome to install. Yielding hydraulic 
jacks, while easier to install and more 
functional, are extremely heavy. With 
this in mind, the Bureau of Mines studied 
ways to reduce the weight of contemporary 
hydraulic supports without sacrificing 
performance. 

In 1979 a contract was awarded to de- 
sign and test a lightweight hydraulic 
mine support. Laboratory testing indi- 
cated some changes were needed in the 
basic design selected. After these modi- 
fications were made, 33 units were sent 
to the field at three different loca- 
tions. These units were designed for use 
in a 6- to 8-ft seam, have a 22-ton ca- 
pacity, are fully self-contained, provide 
a 5- to 7-ton roof preload, and yield to 



overload. The total weight of each unit 
is 55 lb, compared with 110 lb for a com- 
mercial steel unit. All three of the 
Western coal mines were extremely pleased 
with the units and requested to keep 
using them for as long as possible. Dur- 
ing the field testing some problems did 
occur, such as corrosion on the piston 
surfaces , weak pump handles , and short 
life span of some internal seals. All of 
these problems were corrected. 

During the field tests, the need became 
apparent for units that would function in 
seam heights other than 6 to 8 ft. 
Units of the same configuration and ca- 
pacity were built and tested for seam 
heights in the 4.5- to 6-ft, 8- to 10-ft, 
and 10- to 12-ft ranges. After the test- 
ing proved successful, arrangements were 
made to install 10 of the 10- to 12-ft 
units in a mine for field testing. The 
results have not been completed at this 
time. 



INTRODUCTION 



The installing of ten^)orary roof sup- 
port Is an integral part of underground 
coal mining. Historically, wooden posts, 
metal screw jacks, and, more recently, 
telescoping hydraulic supports are used 
for this purpose. Wooden posts are time- 
consuming and cumbersome to install, not 
consistent in size and strength, and also 
represent a sizable constant cost. Screw 
jacks, while easier to install, have no 
yield capability and are prohibitively 
heavy where high strength supports are 
required. The all-hydraulic, telescoping 
cylinder concept remains the best choice 
for reusable temporary support. 

There are now on the market various 
hydraulic supports made of steel tubing 
and steel components that perform very 

^Mining engineer, Spokane Research Cen- 
ter, Bureau of Mines, Spokane, WA. 



well as a temporary support. Unfortu- 
nately, due to increasing support-load 
requirements and the thicker coal seams 
found in the West, these steel supports 
are becoming extremely heavy and cumber- 
some to use. A support rated at 22 tons 
and used in a 6- to 8-ft coal seam weighs 
in excess of 110 lb. In addition, the 
same type support designed for use in a 
12-ft seam weighs approximately 300 lb. 

With this in mind, the Bureau of Mines 
funded a contract to look at state- 
of-the-art technology in lightweight hy- 
draulic cylinder design in an effort to 
construct a high strength-to-weight ratio 
yielding temporary roof support. 

This work was started in 1978 and com- 
pleted in February 1982. This paper 
deals with the evolution of the project 
through the contract and in-house phases. 



130 



CONCEPT EVALUATION AND DETAILED DESIGN 



The contract initially called for exam- 
ination of previous work in the area of 
lightweight supports, as well as existing 
commercially available materials and 
products. The review identified the fol- 
lowing major desirable designs features: 

1. A 22-ton capacity. 

2. A goal weight of 50 lb. 

3. A prop with controlled yielding at 
overload. 

4. A fully self-contained unit (i.e., 
integral reservoir and pressurization 
unit). 



5. A remotely recoverable prop. 

6. A prop easily and quickly 
installed. 

Two concepts selected for detailed design 
incorporated hydraulic mechanisms that 
would yield at overload and be remotely 
recoverable. Of the two designs (totally 
hydraulic and hydromechanical) , the all- 
hydraulic version was selected for proto- 
type production and testing (fig. 1). 
Although the hydromechanical support was 
lighter in weight, it was rejected be- 
cause of its mechanical complexity, cost, 
and complexity of installation in a min- 
ing situation. 



FABRICATION AND TESTING OF LIGHTWEIGHT SUPPORTS 



FABRICATION 

Four prototype models of the all- 
hydraulic lightweight support were fabri- 
cated to verify the design. Based on 
value and manufacturing engineering con- 
siderations, some design changes were 
made prior to release for fabrication. 
Additional changes were found to be 




FIGURE 1. - Six- to eight-hydraulic support. 



necessary during assembly and functional 
testing. Chief among these were modify- 
ing the pump block assembly and redesign- 
ing the main cylinder. 

For structural testing, all four proto- 
type units were modified. The four were 
laboratory tested to ensure functional 
stability and correctness of design. Two 
props were tested to ultimate load fail- 
ure to verify column ultimate strength 
calculations. 

Following the completion of successful 
functional and structural testing of the 
models, a production lot of 40 additional 
units was constructed (fig. 2). It was 
intended that 10 of these supports would 
be provided to each of four different 
mines for a 6-month demonstration and 
evaluation period. 

Following the final assembly, each of 
the production units was subjected to 
operational testing. This consisted of 
pressure testing the internal preload 
limit valve, the 22-ton yield bypass 
valve, and preloading the unit in a test 
frame overnight to detect valve or seal 
leaks. 

An independent testing program was per- 
formed by the Bureau to corroborate the 



131 



Manual pump 
handle 



^k 



Remote retrieval 
lanyard ring 



[J 



-Hydraulic fluid 
reservoir 



Pressure release 
valve 



r-1 



-Carrying handle 
bracket 



1^ 



Telescoping hydraulic 
piston 



FIGURE 2. - Components of the hydraulic support. 

contractors' findings. During these 
tests, one of the reservoir tube welds 
cracked and failed. Subsequent analysis 
by both the Bureau and FMC Corp. 's Mate- 
rials Laboratory revealed that the weld 
was faulty. Other faulty welds were 
detected through radiographic analysis, 
so it was decided to reweld and re- 
heat-treat all of the reservoir tube 
assemblies. 

UNDERGROUND TESTING 



One of the test sites in eastern Utah 
incorporated the supports into a continu- 
ous mining development section. The min- 
ers used the supports continuously, ex- 
cept for the 3-month-long United Mine 
Workers strike (six of the supports were 
left, fully loaded, across an entry for 3 
months with no failures) , and felt the 
lightweight supports were superior to 
what they had been using as far as hand- 
ling and setting were concerned. 

The supports were removed from the mine 
in November 1981 for repair. All of the 
handles were replaced with steel ones. 
Eight of the ten supports were rebuilt 
with cannibalized parts from two supports 
that had been hit by mining equipment and 
were beyond repair. All of the work was 
done locally and sent back to the mine 
for further use. The units were rebuilt 
again in June 1982. The six remaining 
units were then used in a continuously 
advancing development section as part of 
a prop and beam temporary support system. 
These were used until August 1982 when 
they were pulled from the mine for over- 
haul. Due to lack of production, it was 
decided to send the units back to the 
Bureau for rebuilding and inspection. 

Another site, located in western Colo- 
rado, did not install the supports until 
May 1981. Eight units were being used as 
part of their longwall tailgate support 
plan. As of April 1982, the only prob- 
lems encountered were one broken pump 
handle and one sticking release valve. 
After that date the longwall was shut 
down and the supports were used in vari- 
ous applications such as setting brattice 
curtain lines, setting chain-link fence 
at pillar ribs , and in longwall panel 
development work. 



During the latter part of the fabrica- 
tion phase, underground coal mines were 
contacted to solicit interest in partici- 
pation in the demonstration of the light- 
weight hydraulic supports. Three mines, 
two in Colorado and one in Utah, were 
selected to receive the props for under- 
ground testing. Each mine was given 10 
supports to use under actual mining 
conditions. 



The supports were removed from the mine 
in November 1982 for overhaul. In March 
1983, the units were again sent under- 
ground for use in a development section 
and also to be used as additional support 
around a drilling operation for a methane 
drainage project. 

At the third test site, located outside 
Grand Junction, CO, the supports were 



132 



used in development work. In order to 
work in the 8-1/2-ft coal seam, a 2-ft 
steel extension was added to the support. 

The supports were used in slow develop- 
ment work from April until November 1981, 



at which time most of the units needed 
some repair. Problem areas were failure 
of the main seal , sticking pressure re- 
lease valve, and pressure pump piston. 
These supports were sent back to the Bu- 
reau for inspection and rebuilding. 



OVERALL TEST CONCLUSIONS 



In the three test mines, all comments 
about the test were positive. In all 
cases , both miners and management liked 
the supports and wanted to continue to 
use them. Although several areas of 
weakness became evident, they were not 



major problems. It should be noted that 
although the duty-cycle was only about 6 
months , the problems exhibited were con- 
sistent with those props currently on the 
market. 



CURRENT STATUS 



All three test mines have expressed a 
desire to continue using the supports in 
their mining sequence. These mines and 
various other mines have requested infor- 
mation about commercial availability of 
the props. 

A project to test the various lengths 
of lightweight supports is being per- 
formed by the Bureau, since the various 
lengths require slightly different de- 
signs. The first step in this project 
was to have prototype supports built 
using the updated shop drawing package. 
A contract was let to fabricate light- 
weight supports in the 4.5- to 6-ft, 8- 
to 10-ft, and 10- to 12-ft range. These 
supports were received in February 1983. 
A structural testing program to confirm 
design calculations was performed in late 
June 1983. At the same time, modifica- 
tions are being made to the support, 
based on field test results, that will 
increase the duty-cycle and operating 
ability. Some of these changes included: 
redesign of the top end of the cap for 



more bearing surface and to facilitate 
nailing a cap piece on to the longer sup- 
ports prior to setting the unit. Also, a 
redesign of the lanyard ring assembly 
will be performed along with plating the 
pump pressure piston and pressure release 
piston. Adding 3 in to the pump handle 
will help in applying the preload pres- 
sure. The use of various seals in the 
longer units will be investigated. An 
aluminum extension has been fabricated 
and is due for field testing in June 
1983. Three units were fitted with sight 
pressure gauges and sent to Alaska to be 
used during a rock -mechanics project in a 
permafrost gold mine. Six other 6- to 
8-ft units were fitted with pressure 
transducers and will be used in a retreat 
coal mining sequence to monitor loading 
in a breaker prop row. 

Several of the longer supports will be 
fabricated and tested. Evaluations will 
be completed and results published in the 
near future. 



133 



MOBILE ROOF SUPPORT AND APPLICATIONS IN RETREAT MINING 
By Robert R. Thompsoni 



ABSTRACT 



Retreat pillar mining is highly pro- 
ductive, but dangerous. The primary dan- 
ger during pillar removal is premature 
caving of the roof. The Bureau of Mines 
has developed a remotely operated machine 
that will place and retrieve temporary 
roof support. The prototype machine 
worked well but had several problems, the 



primary one being tramming. Two second- 
generation machines were built under a 
cost-sharing program with a Utah coal 
mine and a mining equipment con^jany. The 
machine carries four 50-ton jacks and is 
remotely controlled by radio. The ma- 
chines are presently being tested under- 
ground in a Utah coal mine. 



INTRODUCTION 



Retreat pillar mining is highly pro- 
ductive because supply, haulage, ventila- 
tion, and power systems are established, 
and there is also the advantage of the 
knowledge gained during development such 
as roof and ground behavior and hydro- 
logic factors. The primary danger during 
pillar removal is premature caving of the 
roof. 

The roof must cave in a predictable and 
dependable manner to prevent inducing 
excessive abutment loads in adjacent pil- 
lars, which can result in rib bursts. 



floor heave, or crushed pillars. The 
safety of the miners is dependent on 
successfully controlling the roof. Roof 
support during retreat is usually ob- 
tained by setting posts, cribs, hydraulic 
props, roof bolts, or a combination of 
these devices. They are set manually by 
a miner working in a hazardous area. 
Many of these devices can never be re- 
covered and thus become part of the cost 
of extracting coal. The problem is how 
to set and retrieve these roof supports 
safely. 



MOBILE ROOF SUPPORT SYSTEM 



The Bureau embarked on a project to 
develop a system that would place and 
retrieve temporary roof supports without 
danger to the operator. In conjunction 
with a contractor, a mobile roof support 
(MRS) machine (fig. 1) was designed, 
built, and field-tested. THE MRS was re- 
motely operated, battery-powered, and 
rubber-tired. It carried four jacks, two 
on the body of the machine, and two at 
the end of hinged arms. The jacks extend 

^Research structural engineer, Spokane 
Research Center, Bureau of Mines, Spo- 
kane, WA. 



to form columns between the floor and 
roof, each with 30 tons of potential sup- 
port. The jacks were hydraulically 
locked, and the load distributed to three 
points on each jack, without loading the 
machine chassis. 

The MRS was tested underground in an 
Illinois coal mine. The prototype ma- 
chine worked well but had several prob- 
lems, the primary one being tramming over 
the soft floor. Tramming was slow and, 
at times, the tires became buried and 
machine had to be pulled out. 



134 




FIGURE 1. - First-generation mobile roof support. 



SECOND-GENERATION MOBILE ROOF SUPPORT MACHINE 



Results of the field test were encour- 
aging. The concept had acceptance by 
both mine management and the miners. 
Several industry personnel witnessed the 
field trials and felt the MRS would im- 
prove the safety and productivity of 
their retreat mining sections. With this 
encouragement , the Bureau decided that a 
second-generation machine should be 
built, correcting the problems of the 
prototype. 

A Utah coal mining company offered its 
mine as a test site. It also offered to 
work with the Bureau during the design 



and fabrication of the second-generation 
machine and to share some of its costs. 
This operation removed 12 ft of a 15-ft 
coal seam and had a four-member support 
crew setting posts. This time-consuming 
and hazardous task, at times, occurred 
under unsupported roof. 

A mining equipment company became in- 
terested in producing the MRS. It 
offered to help cost-share by supplying 
some of the parts for the new machines. 
It also offered to assist during the 
design and fabrication. 



135 




FIGURE 2. - Second-generation mobile roof support. 



BASIC MACHINE REQUIREMENTS 



The Bureau embarked on a research pro- 
gram to design, fabricate, and field-test 
two second-generation MRS's (fig. 2). 
During the first 3 months, participants 
met monthly. During this preliminary 
design phase, all participants agreed 
that the machine in the tram mode should 
be 8 ft wide and 10 ft long, or less, 
with at least a 12-in ground clearance. 
Also, it should be of rugged construction 
with towing hooks on both ends and 
mounted on independently controlled 
crawlers that exerted 20 psi, or less, 
ground pressure. The machine should be 



powered by a 40-hp, 460-V ac permissible 
motor from a 260-ft reeled trailing cable 
and transmitter remote controls. Re- 
versible variable tram speed of at least 
80 ft/min on 20 pet grade, and free- 
wheeling for emergency towing, were to be 
included. Front and back dozer blades of 
12-in range are required. 

The machine was to carry two chassis 
mounted and two swing-arm jacks of 7- to 
15-ft working height. The jacks were to 
have a 3- to 8-ton installation and 50- 
ton maximum loading capabilities. The 



136 



swing jacks are to form a breaker row conditions, the jacks will be capable of 



with 6-ft separation between chassis 
jacks and have a visual load indicator. 
In case of heavy ground, the ability to 
remote-jettison either or both swing 
jacks is incorporated. In order to 
attain rapid egress under bad roof 



retraction, so as to provide 1 ft of 
ground clearance and 1 ft of roof clear- 
ance in 30 s. The mine requested that 
the machines be remotely controlled by 
radio. 



UNDERGROUND TESTING 



After the machines are fabricated, they 
are being shipped to the mine to be 
tested for a period of 6 months. Figure 
3 shows the test mine's ground control 
plan during full-pillar extraction. It 
required setting 24 posts for pulling 
each fender. The use of the two machines 
will eliminate the requirement for set- 
ting these posts. Figure 4 shows the 
planned sequential operation with the 
MRS. Again, the machine will be moved 
and set remotely, thus eliminating expo- 
sure of the miners setting the posts by 
hand. Yet to be determined is the 



possible use of some posts to act as 
"squealers" or warning devices. These 
posts, if required, would be set after 
the MRS is in place and supporting the 
roof. 

The mine is expected to use the ma- 
chines for a period of up to 10 jrr , thus 
providing long-term testing. The pro- 
jected mass-produced costs of the ma- 
chines, with tethered remote-control 
rather than radio, are estimated at 
$125,000. The radio remote-control is 
expensive and is considered as an extra. 



• • • • 

• • • • 







FIGURE 3. - Ground control plan during full 
pillar extraction. 




FIGURE 4. - Sequential operation with machines. 



137 



ROOM-AND-PILLAR RETREAT MINING 



The Bureau published a manual for the 
coal industry, 2 which is to provide mine 
managers and engineers with: 

1. Assistance in making decisions to 
retreat mine and in selecting the best 
mining technique for their specific 
condition. 



3. Information to develop a section 
foreman's handbook on retreat mining 
safety and operation. 

Copies may be obtained from the Su- 
perintendent of Documents, U.S. Govern- 
ment Printing Office, Washington, DC 
20402. 



2. Information on efficient retreat 
mining design. 

SUMMARY 



The Bureau has des^-gned and built a 
support system for retreat mining that 
can be set and retrieved remotely. The 

^Kauffman, P. W., S. A. Hawkins, and 

R. R. Thompson. Room and Pillar Retreat 

Mining. A Manual for the Coal Industry. 

BuMines IC 8849, 1981, 228 pp. 



system is now being tested in a Utah coal 
mine. This system provides added safety 
for the miner, by eliminating the need to 
work in a hazardous area setting posts, 
cribs, or hydraulic props. The MRS will 
also increase productivity, since the 
number of manual support setting opera- 
tions has been decreased. 



138 



INORGANIC GROUTS FOR ROOF BOLTING 
By Jack E. Fraley^ 



ABSTRACT 



The Bureau of Mines investigated rapid- 
hardening material substitutes for the 
resin used in mine roof bolts. Gypsum 
plasters (CaS04 • I/2H2O) were selected be- 
cause they have high early strength while 
being readily available and inexpensive. 



Gypsum plaster-water capsule cartridges 
provide a substitute for resin car- 
tridges. Gypsum plaster holds promise 
for injection in roof bolt holes as a 
premixed slurry because of improved oper- 
ator safety and greater economy. 



INTRODUCTION 



Fully grouted resin bolts are a rela- 
tively new phenomenon to the mining in- 
dustry. In 1972, the advantages of resin 
bolts became apparent, and by 1980, an 
estimated 20 million of these bolts were 
installed. Since resin bolts are more 
costly than mechanical bolts, their su- 
perior performance is illustrated by the 
large increase in their usage. 

The price of resin doubled between 1973 
and 1975. Since resin is petroleum- 
derived, its cost and future supply are 
uncertain. Because resin cartridges are 
flammable, they are a potential under- 
ground fire hazard. 

To overcome these resin disadvantages, 
the Bureau of Mines started to investi- 
gate rapid-hardening material substi- 
tutes. Gypsum plasters (CaS04 •I/2H2O) 
were selected because they have high 
early strength while being readily avail- 
able and inexpensive. To achieve desired 
rapid hardening, the plaster is acceler- 
ated by adding 1 pet K2SO4 (by weight of 
dry plaster). 

^Chemical engineer, Spokane Research 
Center, Bureau of Mines, Spokane, WA. 



JV^ 



-Cartridge 



-Voids (air) 



-Cement 



Microcapsules 
of water 



-Grout 



-Rebar 



ABC 

FIGURE 1. - Gypsum-plaster, water-capsule bolt. 



139 



WATER-CAPSULE CARTRIDGES 



A gypsum-plaster, water-capsule car- 
tridge was made, as shown in figure 1. A 
cartridge (packaged similarly to resin 
cartridges) (fig. lA) is inserted into a 
drilled hole. The cartridge wrapper is 
filled with accelerated gypsum plaster 
and water capsules, but also contains air 
as void spaces between the fine gypsum 
particles, as shown in figure IB^. During 
rebar insertion (fig. 1£) , the water cap- 
sules rupture, releasing the water, which 
mixes with the plaster to form hardened 
gypsum. 

Figure 2 shows each component in the 
system. The plaster is on the upper 



left, and the water capsules on the upper 
right. A cartridge is in the center, 
while a short length of rebar is at the 
bottom. 

The water capsules appear in figure 3 
alongside a penny, so their size can be 
noted. Typically, the capsule diameters 
are 1,800 pm (0.071 in). The water cap- 
sules are a modified wax shell surround- 
ing water (encapsulated water). They 
contain over 60 wt pet water; and to be 
of adequate quality, they must retain the 
water and be durable enough to withstand 
normal handling during cartridge produc- 
tion and installation. 




FIGURE 2. - Components of the gypsum-plaster, water-capsule bolt. 



140 




FIGURE 3. - Water capsules. 



During installation, the cartridges are 
manually inserted, as shown in figure 4. 
The plaster-water mixing is shown in fig- 
ure 5. The water capsules contain blue 
dye, which is visible as the rebar is 
inserted through a cartridge inside a 
clear plastic tube. 

The installation procedure for the 
inorganic cartridges is similar to that 
for resin cartridges, except that less 
rotation for mixing is required. After 
the cartridges are inserted in a drilled 
hole, the bolt is inserted part way while 
being rotated. A wrench is placed on the 
bolting machine rotation chuck, which 
allows complete insertion with the avail- 
able vertical movement. After the wrench 



is in place, the bolt is rotated during 
the remainder of the insertion. Extended 
spinning, with the bolt fully inserted, 
is not required. As the bolt is in- 
serted, pressure builds within the hole 
and ruptures the water capsules. The wet 
mix that is formed is stiff enough so the 
bolt remains in place, allowing immediate 
lowering of the bolter head. 

Ninety percent of the gypsum strength 
is developed in less than 10 min. Pull 
strengths of the bolts (4-ft, Grade 50, 
No. 6 rebar) done in concrete blocks 
exceeded the bolt yield point which 
is over 22,000 lb. Table 1 shows the 
pull strengths 10 to 13 min after 
installation. 



141 




FIGURE 4. - Manual insertion of cartridges. 



TABLE 1. - Pull strength of gypsum-water 
capsule bolts 



Bolt 


Pull, lb 


Time , mln 


2 


26,150 
22,400 
21,400 
25,200 
22,400 
21,400 
21,400 
24,300 
22,400 


10 


3 


13 


4 


10 


5 


10 


6 


10 


7 


10 


9 


11 


11 


12 


17 


10 



When the bolt is thrust into a confined 
cartridge, the pressure has an effect on 
subsequent pull strength. Pressure as 
high as 5,500 psi has been measured dur- 
ing bolt insertion in 4-ft holes. Four 
pull test samples were made by pouring 
concrete in 4-in pipes that were 4 ft 
long. After installation of 4-ft bolts 
in the samples , the samples were cut into 
eight 6-in sections. Each section had an 
extension rod attached to the bolt for 
pull testing. The average pull strength 
for the four 6-in sections closest to the 



142 



bolt head (sections A-D) , as well as that 
for the four 6-in sections farthest from 
the bolt head (sections E-H) , is given in 
table 2. The pull strengths were higher 
in the half of the bolt furthest up the 
hole, where pressures were higher, owing 
to increased cartridge confinement. 

TABLE 2. - Average pull strength of 6-in 
bolt section, pounds 



Bolt 



Section A-D 
average 



Section E-H 
average 



1. 
8., 

12, 
16, 



8,119 

6,450 

11,156 

14,725 



8,756 

6,963 

17,538 

14.794 



Bolts were installed in foamed concrete 
to measure the effect of the weaker rock 
on pull strength. One bolt gave 9,000 lb 
and another 15,700 lb pull. More spin- 
ning during bolt insertion appears to en- 
large the hole in weaker materials like 
foamed concrete. The smaller amount of 
required spinning in installing the 
gypsum-water capsule bolt may become an 
advantage in weaker rock. 

The cartridges are made to have a 
water-to-plaster weight ratio of 0.30 to 
0.35. Less water gives Increased 
strength; however, rebar insertion is 
more difficult, so the 0.30 weight ratio 
is the approximate lower limit for re- 
peatable rebar installation. 

An additive that increases the fluidity 
of freshly mixed grout at any water- 
to-plaster ratio makes the mixing and re- 
bar insertion easier and reduces bolt in- 
sertion time. The additive allows rebar 
insertion at water-to-plaster ratios as 
low as 0.21. This provides a greater 
margin for successful underground instal- 
lation with variations in the cartridges 
and operator-equipment techniques. By 
lowering the quantity of water as water 
capsules, cost savings of a couple cents 
per cartridge are possible. 



FIGURE 5. - Mixing of gypsum plaster and 
water capsules. 



A research contract was issued to de- 
velop the packaging technology for com- 
mercial production of the cartridges. 
An encapsulation system to provide the 



143 



water capsules was developed along with 
the equipment to produce 20 cartridges 
per minute. The contractor needs private 



venture capital to develop a complete 
production plant and marketing system for 
the cartridges. 



WATER TUBE CARTRIDGES 



Cartridges that replace the water cap- 
sules with a tube of water are being 
studied. Figure 6 shows two methods of 
storing the water within the cartridge. 
Panel A shows a continuous-length water 
tube inside the cartridge. As the rebar 
begins to push on the cartridge, the com- 
pression shortens the cartridge length. 
Flexible water tubes like lay-flat tubing 
are thought to bend as the cartridge com- 
presses, rather than rupture and release 
the water. The result is hard rebar 
insertion due to poorly mixed plaster. 

A semirigid tube that retains its shape 
releases its water sooner to enhance mix- 
ing and ease rebar insertion. Several 
tube materials have been investigated to 
find one capable of retaining water, 
withstanding normal handling during 



V ////////////y////// 



Gypsum 





Water tube 



- f^Z^QOO 



B 



/ A Water package 



C J Gypsum package 



FIGURE 6. - Water tube cartridges. 



cartridge preparation and installation, 
and breaking as soon as the cartridge is 
compressed. 

Figure 6B^ illustrates alternate pack- 
ages of water and plaster within the car- 
tridge wrapper. As with the continuous- 
length water tube, best rebar insertion 
is obtained if the water is released as 
the cartridge is first compressed. 

To reduce the loss of fluid material 
from the hole during rebar insertion, a 
plastic cap can be inserted over the 
rebar , which plugs the hole as the rebar 
installation progresses (fig. 7). 




r~ji 



Cap 



FIGURE 7, - Cap to reduce loss of fluid from 
hole during rebar insertion. 



144 



Water tube cartridges require more mix- 
ing than water capsule cartridges because 
the water is not dispersed throughout the 
cartridge. To ease water-plaster mixing, 
an additive can be mixed with the plaster 



to increase its fluidity at a given water 
content. Also, the surface tension of 
the water can be reduced to increase its 
ability to wet the plaster. 



SLURRY INJECTOR 



The slurry injector is a bulk injection 
method that lends itself well to remote- 
control operations. The operator can 
work under supported roof during full- 
column bolting. Dry gypsum plaster is 
automatically mixed with water to form a 
slurry, pumped into a delivery hose, and 
injected up the roof -bolt hole, without 
placing any mechanical device in the 
hole. The system uses a twin-screw ex- 
truder (fig. 8) normally used for pro- 
cessing plastics. The geometry of the 
screws makes them self-cleaning. The 
extruder mixes and pumps the grout into a 
20-ft delivery hose attached to a trans- 
fer device. After the grout is in the 
hose, the transfer device inserts a 
plastic "rabbit" or plug behind the 
grout. High-pressure air then drives the 



rabbit and grout through the hose and 
nozzle into the roof -bolt hole. The 
delivery hose is cleaned by the rabbit. 
The nozzle is positioned beneath the hole 
by a linkage mounted on the bolter drill 
head. 

To operate the system, the hopper is 
filled with gypsum plaster and a tank is 
filled with water. A control knob se- 
lects the proper grout volume and water- 
to-plaster weight ratio for the size hole 
being drilled. After the hole is 
drilled, the nozzle is positioned under 
the hole and the rabbit inserted into the 
transfer device. The machine is switched 
on and all operations through complete 
bolt insertion are then automatic. 



Transfer device 



Transmission 




Delivery hose 



screws 



Knife 
valve 



Hydraulic 
motor 



FIGURE 8. • Slurry injector components. 



CONCLUSIONS 



145 



1. Gypsum plaster-water capsule car- 
tridges provide adequate bonding of No. 6 
rebar in 4-ft holes. 



4. Capital is necessary for commercial 
production of the gypsum plaster-water 
capsule cartridges. 



2. The gypsum plaster-water capsule 
cartridges provide a substitute for resin 
cartridges. 



5. Water tube cartridges should be 
further investigated, as they may be eas- 
ier to manufacture and more economical. 



3. The gypsum plaster-water capsule 
cartridges may be an advantage in softer 
rock where more spinning disturbs the in- 
tegrity of the hole. 



6. The slurry injector is a promising 
concept because of possible improvements 
in operator safety and econoiny of bolt 
installations. 



146 



RESEARCH, DEVELOPMENT, AND USE OF STEEL-FIBER-REINFORCED 
CONCRETE CRIBBING FOR MINE ROOF SUPPORT 

By Thomas W. Smelser^ and Dale A. Didcoct^ 



ABSTRACT 



Through the combined efforts of the health and safety conditions with a 



Bureau of Mines and private industry, 
steel-fiber-reinforced concrete crib 
block to improve coal mine safety in the 
area of roof control has been developed 
and is currently in use commercially. 
The development objective was to improve 



stiffer, stronger, nonflammable roof sup- 
port at a competitive cost. During the 
time this method of roof control has been 
in use in the mining industry, it has 
proven to have definite functional and 
economic advantages. 



INTRODUCTION 



In 1975, the Bureau initiated develop- 
mental research in the area of a steel- 
fiber-reinforced concrete crib block as 
an alternative to the use of wood cribs. 
Aside from the obvious disadvantages of 
low stiffness and strength, deterioration 
from chemical and bacteriological attack, 
methane liberation, and f lammability , 
wood cribs are subject to variables in 
cost and availability of suitable vari- 
eties in many areas of the country. In 
the Eastern States, a reliable source of 
pressure-treated wood is often difficult 
to find, and the scarcity of any type of 
wood in some Western States makes it ex- 
tremely expensive. Concrete offered the 
lowest cost support and appeared to be 
the best choice for replacing wood sup- 
ports if a solution could be found for 
its poor failure characteristics. The 
addition of steel fibers to the concrete 
would greatly improve the safety factor 
by eliminating the possibility of sudden 
brittle failure. 

Starting in 1976, a project was con- 
ducted at the Bureau's Spokane (WA) Re- 
search Center, under the direction of 
G. L. Anderson, Research Structural En- 
gineer, and Thomas W. Smelser , Super- 
visory Mechanical Engineer. The many 

^Supervisory mechanical engineer, Spo- 
kane Research Center, Bureau of Mines, 
Spokane , WA . 

^vice president. Underground Technology 
Div., Burrell Group, Morristown, TN. 



types of fibers that were considered in 
the initial testing included glass, as- 
bestos, aramid, nylon, carbon, polypropo- 
lene, and steel. Except for steel, the 
aforementioned fibers were each elimi- 
nated owing to various problems with mix- 
ing characteristics, economy, health haz- 
ards, strength, etc. Steel fibers also 
appeared to be the lowest cost solution, 
as compared with the use of steel- 
reinforcing bars or steel wire mesh. The 
steel fiber types considered included 
straight-round, straight-flat, crimped- 
full-length, melt-extracted, deformed- 
full-length, and bent-end. The bent-end 
type fiber was selected, based on per- 
formance and cost considerations. 

Upon completion of the selection of the 
fibers to be used, the next step was to 
select a concrete mix design and deter- 
mine the quantity of fibers for the mix. 
The steel-fiber-reinforced concrete (SFC) 
support members that were developed 
offered a significant improvement in 
stiffness and strength in compression 
compared with wooden cribs, yet they 
avoided the brittle or catastrophic com- 
pressive failure mode of plain concrete. 
Full-scale compression testing showed the 
ultimate compressive strength of SFC 
blocks to be 4,000 psi, compared with 500 

psi for wood. Because eight times more 
wood is necessary to equal the strength 
of the steel-fiber-reinforced concrete 
crib block, the use of SFC greatly in- 
creases the area for movement of 



147 



personnel and equipment and for ventila- 
tion airflow. The 4,000-psi concrete was 
selected as optimum strength by most 
closely approximating the strength of a 
typical roof and floor structure of a 
coal mine. 

Bureau Report of Investigations 8412, 
published in 1980, contains support sys- 
tem design and results of laboratory 
investigation and full scale testing. 3 

A successful installation of steel- 
fiber-reinforced concrete cribs at 
Kaiser's Sunnyside Mine in Utah was made 
in 1976 as part of the single-entry 



longwall demonstration at that mine. The 
wood cribs being used to support the 
single entry were not stiff enough to 
hold the tailgate section of the entry 
open after passage of the first longwall 
face. Although limited, the demonstra- 
tion showed the promise of concrete mine 
support systems and resulted in the 
project, covered by this report, to more 
thoroughly characterize and evaluate 
materials in the laboratory for improved 
support characteristics, safety, and 
economy. A follow-on program field- 
tested the support systems to verify in- 
stalled cost, structural behavior, and 
industry acceptance. 



DEVELOPMENT OF MANUFACTURING PROCESS 



The prototype blocks made by the Bureau 
in the research and testing phase were 
manufactured on a small production basis 
and cost approximately $7 per block. 
This cost was prohibitive, and a meth- 
od of mass-producing the steel-fiber- 
reinforced blocks had yet to be devel- 
oped. A major manufacturer of concrete 
block and steel fiber gunite mixes, Bur- 
rell Construction and Supply Co., New 
Kensington, PA, was contacted by the re- 
search team. 

With the Bureau contributing their 
technical specifications, and Burrell 
Construction and Supply Co. providing the 
formulation, production technique, and 

■^Anderson, G. L., and T. W. Smelser. 
Development Testing and Analysis of 
Steel-Fiber-Reinforced Concrete Mine Sup- 
port Members. BuMines RI 8412, 1980, 
38 pp. 



mix methods, an economical SFC block was 
produced in the summer of 1980. 

By summer 1981, Burrell had developed a 
process and method (patent applied for) 
for mixing the concrete and steel fibers 
to produce crib blocks that fulfilled the 
standards set by the Bureau. Many full- 
scale crib tests were performed at Lehigh 
University, Pittsburgh Testing Lab, and 
on the Mine Roof Simulator at the U.S. 
Department of Energy (DOE) in Bruceton, 
PA (figs. 1-4). The results of the DOE 
testing are contained in the DOE Report 
MRS-DR-81-05.4 

^yrd, R. J., and J. L. Thompson. Mine 
Roof Simulator Data Report Steel- 
Fiber-Reinforced Concrete Roof Cribs 
(Three Configurations). U.S. Dep. 
Energy, Min. Equip. Test Facility, MRS- 
DR-81-05, Aug. 1981, 40 pp. 



148 




FIGURE 1. - Solid crib failure mode (DOE tests). 



2,100 




ANALOG DATA 


Text No. 30901 


1 


1 1 1 1 


1 1 


1,800 


- 




- 


1,500 


A 




- 


1,200 


- / 


\ 


- 


900 


- / 


\ 


- 


600 


- / 


\ 


- 


300 


y 


\...^ 


- 




1 1 1 1 


1 1 



0.25 0.50 0.75 1.00 1.25 ISO 1.75 2.00 

Text No. 30902 




0.25 0.50 0.75 1.00 1.25 1.50 1.75 2.00 

VERTICAL DISPLACEMENT, in 

FIGURE 2. - Solid crib-vertical load versus 
displacement (DOE tests). 




FIGURE 3, • Open crib failure mode (DOE tests). 



ANALOG DATA 



Test No. 30903 




0.50 0.75 1.00 1.25 1.50 

VERTICAL DISPLACEMENT, in 



FIGURE 4. - Open crib-vertical load versus 
displacement (DOE tests). 

The resulting mass production technique 
reduced the cost per unit to $2.29 FOB 
plant. The block is presently being man- 
ufactured under licensed agreement with 
Burrell at locations in Pennsylvania, 
Ohio, Virginia, Utah, and Alabama. Other 
block producers are beginning to enter 
the market with similar products. 



149 



UNDERGROUND EXPERIENCE 



Since the original installations in 
1976, at Kaiser's Sunnyside Mine, the 
Bureau has been involved in cooperative 
underground evaluations at several other 
locations: 

Kaiser Steel, York Canyon, NM 

Price River Coal, Helper, UT 

Snowmass Coal, Carbondale, CO 

Bethlehem #131, Van, WV 

U.S. Steel #9, Gary WV 

In all these cases, the cribs were used 
in the tailgate section of the entries on 
longwall panels (fig. 5). Because of the 
successes of the initial installations, 
some of the above are expanding the use 
of SFC to areas where permanent support 



systems are required, such as ventilation 
entries, and to support the base of ven- 
tilation shafts. 

Several other coal companies have 
elected to institute the use of this 
product in longwall entries, long- 
life main entries, bleeder entries, and 
stoppings: 

Eastern Associated Coal Company — WV 

Emery Mining Company — UT 

Trail Mountain Mining Company — UT 

Barnes & Tucker #25 — PA 

Penn Allegheny Coal — PA 

Canterbury Coal — PA 




FIGURE 5. - Concrete cribs supporting tailgate after passage of longwall face. Kaiser's York 
Canyon, NM, mine. 



150 



Carpentertown Coal & Coke — PA 
Scott's Branch Mine — KY 
Martin County Coal — KY 
Texas Gulf, Inc. — WY 
Westmoreland Coal Company — VA 
Helvetia Coal Company — PA 
ARMCO Inc.~WV 
Consolidation Coal Co. — WV & PA 



Island Creek Coal Co. — VA 

Jim Walters Resources Inc. — AL 

Kit Energy — WV 

In almost all applications of the tail- 
gate entries for longwall, the SFC has 
not only provided added safety, better 
airflow, and more area for movement, but 
has also proven cost-effective. For 
example, in a mine using a double row of 
wood cribs set on 5-f t centers , they are 
now installing SFC cribs in a single row 
on 7-ft centers (figs. 6-7). 




FIGURE 6. - Longwall tailgate entry with double row of wood cribbing. West Virginia mine. 



151 




FIGURE 7c, - Longwall tailgate entry with single row of concrete cribbingo West Virginia mine,, 



COST ANALYSIS 



Following is an example of a cost com- 
parison for this West Virginia mine: 

FIBERCRIB versus Wooden Cribs 

(Based on 84-in height — 1,000-ft 
advancement 

Wooden Cribs - (6 by 8 by 30 in) — Double 
Row on 5-ft centers — 
28 blocks per crib 
@ $2.39 FOB Mine = $66.92 
per crib 

400 Cribs (g $66.92 = $26.768.00 



FIBERCRIB Cribs 



(3-5/8 by 7-5/8 by 
23 in) —Single Row 
on 7-ft centers — 
46 blocks per crib 
(a $2.88 FOB Mine 
= $132.48 per crib 



142 Cribs @ $132.48 = $18,812.00 



Labor Costs - The labor cost to build 
each crib is approximately equal. Note 
that 400 wood cribs were required in this 
example, as opposed to 142 FIBERCRIB 
cribs, resulting in a labor cost saving 
of over 60 pet. The total resulting cost 
saving in this example is over 40 pet for 
the FIBERCRIBS. 

In another instance, where the mine was 
setting a single row of wood cribs skin- 
to-skin, they have found the use of SFC 
cribs on 6-ft centers to be cost- 
effective. 

In an Alabama mine, a longwall tail- 
gate, originally supported with a single 
row of wooden cribbing 9 ft on center, is 
being successfully supported with a 
single row of SFC cribs 15 ft on center. 
This will result in a total savings of 
over 55 pet for labor and materials ex- 
pended for support of this tailgate 
(figs. 8-9). 



152 




FIGURE 8. - Tailgate supported with concrete cribbing ahead of longwall face. Alabama mine. 



Many of these mines have also found the 
SFC cribs to be efficient in areas where 
longevity is a factor and, although the 
cost may be initially higher in some 
cases, the longer life of the SFC crib- 
bing will make them more economical in 
the long run. The industry has found 
that in the case of long-life entries, 
the SFC cribbing is an advantage because 
of the inherent qualities of stiffness, 
strength, nonshrinking, nonrotting, and 
nonflammability. Replacement costs on 
wood cribbing in these applications have 
proven to be very high. In any area 
where a wood crib would need replacing 
because of rot and/or failure, the SFC 
block would eventually prove to be more 
economical. 



In most applications it is possible to 
use fewer SFC cribs than wood cribs in 
order to achieve the same support. A 
Pennsylvania mine, with their belt and 
track lines running parallel, was using 
wood posts skin-to-skin and, in some 
cases, two to three deep to support the 
roof. They are now building solid SFC 
cribs on 6-ft centers with an "I" beam 
from crib to crib. This has given them 
much greater support and they now have 
access between beltline and track. This 
particular crib installation was in- 
spected by representatives of the Office 
of Deep Mine Safety, Pennsylvania Depart- 
ment of Environmental Resources. The two 
inspectors recommended, "An approval be 
granted to Burrell Construction and 



153 




fT^ 




^»*i.^*»" 




FIGURE 9, = Tailgate supported with concrete cribbing after passage of longwall face.. Alabama mine. 



Supply Company for use of FIBERCRIB in 
bituminous underground coal mines in 
Pennsylvania, providing FIBERCRIB blocks 
strictly comply with the manufacturer's 
specifications" (fig. 10). 

To date, a combined total of about 20 
miles of longwall tailgate entry, main 
entry, and bleeder entry is being 



supported with SFC cribbing in the major 
U.S. coal mining regions. SFC cribbing 
has also been used to build stoppings and 
overcasts where normal cinder block had 
proved ineffective due to crushing. The 
SFC block is approximately 2-1/2 times 
stronger than a regular cinder block and 
is not subject to brittle failure. 



154 




FIGURE lOo " Solid concrete cribs placed 6 ft on center supporting a main haulage entryo 
Pennsylvania mine. 



INSTALLATION PROCEDURES 



The procedure for installation of SFC 
crib blocks, suggested by the Bureau, is 
as follows: 

CONCRETE CRIBS (FIBERCRIB) 

1. Prepare floor area level and flat, 
large enough for 23- by 23-in, 15- by 23- 
in, or 15- by 15-in crib, as required. 

2. Place first layer of crib blocks 
and check for level with hand level. 

3. Stack block in open or solid con- 
figuration according to plan. 

4. While stacking, keep blocks 
straight, square, and plumb and remove 
dirt, etc., from each layer before plac- 
ing next layer . 

5. Top of crib must be finished with 
wood plank or beam and wedged tightly 



with wood wedges. The total thickness of 
wood including wedges must be a minimum 
of 1 in for each foot of crib height (4 
in for 4-ft crib, 6 in for 6-ft crib, 
etc.) (fig. 11). 

CONCRETE STOPPINGS USING 
FIBERCRIB BLOCKS 

1. Follow above procedure except ex- 
cavate floor for stopping dimensions. 

2. Stack block flat and in overlapping 
fashion as brick is typically laid. 

3. Finish top of stopping with the 
least amount of wood wedging possible or 
fill with mortar mix. 

4. Seal one side of stopping with 
suitable mortar or sealant. 



155 




FIGURE 11. - Example of insufficient wood at top of crib. Point loading. 



SUMMARY AND CONCLUSIONS 



There is no question of the strength 
and longevity of the steel-fiber- 
reinforced concrete cribbing and its suc- 
cess in present installations. Although 
wood has been the traditional material in 
use for roof support in the mining in- 
dustry, the problems of supply and cost 
are increasing rapidly to the point where 
a viable alternative must be found. Con- 
crete products are readily available in 
all parts of the country, and, indeed, 
the world. The steel-fiber-reinforced 
concrete cribs are proving to be the 
answer to many roof control problems 



experienced in the past and are 
economical alternative to wood. 



the most 



One industry source in the field of 
mine safety asserted: "The new cribs are 
much safer due to the inherent character- 
istics of steel reinforced concrete. We 
have had no failures, and we have long- 
term installations in which we have con- 
fidence that these cribs are a reliable 
source of roof support." Some mines have 
already decided that, in their opera- 
tions, wood cribbing is a thing of the 
past. 



INT.-BU.OF MINES, PGH., PA. 27440 



H 23b 84 



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HECKMAN 

BINDERY INC. 

^^ NOV 84 

^fey N. MANCHESTER, 
^=^ INDIANA 46962 



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